Pt amir hajar kilsi cablebolting in-underground_mines

Copyright © 1996 D. Jean Hutchinson, Mark S. Diederichs and
Geomechanics Research Centre
All rights reserved
No part of this book may be reproduced, stored in a retrieval system or
transmitted in any form whatsoever or by any means without written
consent of the authors.
Layout design by Mark S. Diederichs
Published by BiTech Publishers Ltd.
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Canada V7A 5G1
Fax: (604)277-8125
Printing and binding by Friesens
Altona, Manitoba
Printed in Canada First Printing, March 1996
Cover Note:
Modern cablebolt configurations (top to bottom):
plain strand, birdcaged strand, nutcaged strand, bulbed strand
Canadian Cataloguing in Publication Data
Hutchinson, Douglas Jean, 1961-
Cablebolting in underground mines
Includes bibliographical references.
ISBN 0-921095-37-6
1. Mine roof bolting. I. Diederichs, Mark S. (Mark Stephen),
1964-II. Title. TN289.3.H87 1996 622' .28 C96-910214-3
iii
Table of Contents
Foreword . . . . . . . . . . . . . . . . . . . . . . . vii
Preface . . . . . . . . . . . . . . . . . . . . . . . .viii
Project Sponsors . . . . . . . . . . . . . . . . . . . . . ix
Chief Advisors and Contributors . . . . . . . . . . . . . . . . ix
Correspondence . . . . . . . . . . . . . . . . . . . . . ix
Disclaimer . . . . . . . . . . . . . . . . . . . . . . . ix
Acknowledgements . . . . . . . . . . . . . . . . . . . . x
Chapter 1
Introduction: Cablebolting in Underground Hard Rock Mines
1.1 What is a Cablebolt? . . . . . . . . . . . . . . . . 1
1.2 Why Cablebolt? . . . . . . . . . . . . . . . . . 2
1.3 Cablebolt Applications . . . . . . . . . . . . . . . 5
1.4 The Cablebolting Cycle . . . . . . . . . . . . . . . 8
1.4.1 Design . . . . . . . . . . . . . . . . 9
1.4.2 Implementation . . . . . . . . . . . . . 9
1.4.3 Verification . . . . . . . . . . . . . . . 9
1.5 The Cablebolt Toolbox . . . . . . . . . . . . . . . 10
1.6 Cablebolt Function . . . . . . . . . . . . . . . . 12
1.7 Equipment . . . . . . . . . . . . . . . . . . . 14
1.8 Cablebolt Installation Options . . . . . . . . . . . . . 15
1.8.1 The Breather Tube Installation Method . . . . . . 16
1.8.2 The Grout Tube Installation Method . . . . . . . 17
1.8.3 The Retracted Grout Tube Installation Method . . . . 18
1.8.4 The Grout and Insert Installation Method . . . . . 19
1.9 The Cost of Cablebolting . . . . . . . . . . . . . . 20
1.10 A Note about Units . . . . . . . . . . . . . . . . 21
1.11 Useful Definitions . . . . . . . . . . . . . . . . . 22
Chapter 2
Design: Application of Engineering Principles
2.1 Introduction . . . . . . . . . . . . . . . . . . 23
2.1.1 Design Acceptability Criteria . . . . . . . . . 24
2.2 Capacity and Demand . . . . . . . . . . . . . . . 25
2.2.1 Introduction . . . . . . . . . . . . . . 25
2.2.2 Loading Configurations for Cablebolts . . . . . . 28
2.3 The Cablebolt System . . . . . . . . . . . . . . . 34
2.3.1 The Cablebolt Element . . . . . . . . . . . 35
2.3.2 The Cablebolt Array . . . . . . . . . . . . 36
2.4 7-wire Steel Strand . . . . . . . . . . . . . . . . 37
2.4.1 Definitions . . . . . . . . . . . . . . . 37
2.4.2 Strand Construction . . . . . . . . . . . . 39
2.4.3 Strand Performance . . . . . . . . . . . . 42
2.4.4 Strand Capacity Considerations . . . . . . . . 44
2.4.5 Corrosion of Steel Strand . . . . . . . . . . 45
2.5 Grout . . . . . . . . . . . . . . . . . . . . . 51
2.5.1 Composition of Cement Grout . . . . . . . . . 52
2.5.2 Varieties of Portland Cement . . . . . . . . . 53

iv
2.5.3 Care and Quality of Cement and Water . . . . . . 55
2.5.4 Properties of Fresh Cement Paste . . . . . . . . 56
2.5.5 Properties of Hydrated Portland Cement . . . . . . 65
2.5.6 Cement Grout Specifications for Cablebolting . . . . 71
2.5.7 Grout Admixtures . . . . . . . . . . . . . 72
2.6 Load Transfer . . . . . . . . . . . . . . . . . . 76
2.6.1 Bond Strength . . . . . . . . . . . . . . 77
2.6.2 Bond Strength of Plain Strand Cablebolts . . . . . 79
2.6.3 Modified Geometry Strand . . . . . . . . . 101
2.6.4 Debonding . . . . . . . . . . . . . . 104
2.6.5 Double and Multiple Strand . . . . . . . . . 106
2.6.6 Grout and Rock Shear Strength . . . . . . . . 108
2.6.7 Load Transfer and Surface Anchorage . . . . . 110
2.7 Surface Anchorage and Retention . . . . . . . . . . . 111
2.7.1 Plates . . . . . . . . . . . . . . . 111
2.7.2 Surface Anchorage - Barrel and Wedge . . . . . 112
2.8 Shear Loading of Cablebolts . . . . . . . . . . . . 121
2.8.1 Direct Shear . . . . . . . . . . . . . 121
2.8.2 Oblique Loading - Shear . . . . . . . . . . 122
2.8.3 Cablebolt Orientation . . . . . . . . . . . 124
2.9 Cablebolt Strand Alternatives . . . . . . . . . . . . 125
2.9.1 Plain Strand . . . . . . . . . . . . . 126
2.9.2 Epoxy Coated/Encapsulated Strand . . . . . . 128
2.9.3 Swaged/Buttoned Strand . . . . . . . . . . 129
2.9.4 Birdcaged Strand . . . . . . . . . . . . 131
2.9.5 Nutcaged Strand . . . . . . . . . . . . 133
2.9.6 Bulbed Strand . . . . . . . . . . . . . 134
2.9.7 Combination Strand . . . . . . . . . . . 136
2.9.8 Strand Selection . . . . . . . . . . . . 136
2.9.9 Strand Alternatives: Fibreglass Cablebolts . . . . 138
2.10 Installation Configuration . . . . . . . . . . . . . 140
2.10.1 Grout Mix Design Selection . . . . . . . . . 140
2.10.2 Cablebolt Installation Method Selection . . . . . 141
2.10.3 Borehole Diameter Specification . . . . . . . 141
2.11 Selection of Installation Equipment . . . . . . . . . . 149
2.11.1 Drilling Equipment . . . . . . . . . . . 149
2.11.2 Grouting Equipment . . . . . . . . . . . 150
2.11.3 Breather and Grout Tubes . . . . . . . . . 158
2.11.4 Installation Accessories . . . . . . . . . . 159
2.12 Pipe Pumping Test Procedures . . . . . . . . . . . . 164
2.13 Demand . . . . . . . . . . . . . . . . . . . 166
2.13.1 Excavation Response . . . . . . . . . . . 167
2.13.2 Stress - A Brief Introduction . . . . . . . . . 169
2.13.3 Strength . . . . . . . . . . . . . . . 174
2.13.4 Block Size and the Influence of Scale . . . . . . 175
2.14 Rockmass Classification . . . . . . . . . . . . . . 177
2.14.1 Rockmass Classification Components . . . . . . 178
2.14.2 Data Collection . . . . . . . . . . . . 179
2.14.3 Rock Quality Designation, RQD . . . . . . . 182
2.14.4 Rock Mass Rating, RMR . . . . . . . . . . 186
2.14.5 Rock Tunnelling Quality Index, Q . . . . . . . 191
2.14.6 Modified Rock Quality Index, Q' . . . . . . . 197
2.14.7 Comparison of Rockmass Classifications . . . . . 198
v
2.15 Rockmass Properties from Classification Systems . . . . . . 200
2.15.1 Rockmass Strength . . . . . . . . . . . 200
2.15.2 Stiffness: Rockmass Modulus . . . . . . . . 202
2.16 Empirical Design . . . . . . . . . . . . . . . . 206
2.16.1 Rock Quality Designation, RQD . . . . . . . 207
2.16.2 Rock Mass Rating, RMR . . . . . . . . . . 208
2.16.3 Rock Tunnelling Quality Index - Q . . . . . . . 213
2.16.4 Empirical Cablebolt Design - General Limits . . . 217
2.16.5 Empirical Design - Rules of Thumb . . . . . . 218
2.17 Empirical Design of Open Stopes and Support:
Mathews/Potvin Stability Graph Method . . . . . . . . . 221
2.17.1 Modified Stability Number, N' . . . . . . . . 222
2.17.2 Stability Graph Method - Input Parameters . . . . 223
2.17.3 Open Stope Case History Database . . . . . . 230
2.17.4 Semi-Empirical Cablebolt Design Approach . . . . 236
2.17.5 Stability Graph - Examples . . . . . . . . . 243
2.17.6 Stability Graph Method - Limitations . . . . . . 246
2.17.7 Stability Graph - Calibration to Local Conditions . . 248
2.17.8 Parametric Analysis . . . . . . . . . . . 249
2.17.9 Probabilistic Analysis . . . . . . . . . . 251
2.17.10 Dilution and the Stability Graph . . . . . . . 252
2.18 A Mechanistic Toolbox: Customizing the Design . . . . . . 253
2.18.1 Stress Induced Boundary Crushing . . . . . . 254
2.18.2 Stress Shadowing and Relaxation . . . . . . . 255
2.18.3 Limiting Displacement - Reinforcement . . . . . 256
2.18.4 Stress Induced Joint Slip . . . . . . . . . . 258
2.18.5 Dynamic Loading . . . . . . . . . . . . 259
2.18.6 Surface Unloading . . . . . . . . . . . 260
2.18.7 Sliding Wedge . . . . . . . . . . . . . 260
2.18.8 Two-Dimensional Wedge . . . . . . . . . . 261
2.18.9 Three-Dimensional Wedge . . . . . . . . . 262
2.18.10 Stress Induced Buckling: Euler Approach . . . . 263
2.18.11 Drifts and Intersections . . . . . . . . . . 264
2.18.12 Gravity Bending/Buckling:
No-tension Slab - Voussoir Approach . . . . . . 265
2.18.13 Other Applications . . . . . . . . . . . 275
Chapter 3
Implementation: Making the Design Work
3.1 Introduction . . . . . . . . . . . . . . . . . 277
3.2 The Cablebolting Crew . . . . . . . . . . . . . . 279
3.2.1 Crew Tasks . . . . . . . . . . . . . . 279
3.2.2 Crew Composition . . . . . . . . . . . . 280
3.2.3 Crew Training . . . . . . . . . . . . . 280
3.2.4 Crew Payment . . . . . . . . . . . . . 281
3.3 Training . . . . . . . . . . . . . . . . . . . 282
3.3.1 Why Use Cablebolts? . . . . . . . . . . . 284
3.3.2 What is a Cablebolt? . . . . . . . . . . . 284
3.3.3 How are Cablebolts Installed and Checked? . . . . 285
3.3.4 Safety . . . . . . . . . . . . . . . 287
3.3.5 Feedback on Installation Procedures . . . . . . 287
3.4 Communication . . . . . . . . . . . . . . . . 288

vi
3.5 Quality Control Practice . . . . . . . . . . . . . . 290
3.6 Installation . . . . . . . . . . . . . . . . . . 291
3.6.1 Safety Guidelines . . . . . . . . . . . . 293
3.7 Material Purchasing and Handling . . . . . . . . . . . 295
3.7.1 Design Specifications . . . . . . . . . . . 296
3.7.2 Procedure and Safety . . . . . . . . . . . 297
3.8 Cablebolt Borehole Preparation . . . . . . . . . . . . 298
3.8.3 Quality Control . . . . . . . . . . . . 303
3.8.4 Feedback . . . . . . . . . . . . . . 304
3.9 Cablebolt Installation . . . . . . . . . . . . . . . 305
3.9.3 Quality Control . . . . . . . . . . . . 336
3.9.4 Feedback . . . . . . . . . . . . . . 340
3.10 Automated Cablebolting Systems . . . . . . . . . . . 341
3.10.1 Automated System Design Specification . . . . . 341
3.10.2 Automated System Procedure and Safety . . . . . 343
3.10.3 Automated System Quality Control . . . . . . . 345
3.10.4 Automated System Feedback . . . . . . . . 345
3.11 Quality Control Monitoring and Testing . . . . . . . . . 346
3.11.1 Effect of Quality Control on Cablebolt Capacity . . 346
3.11.2 Checking Quality Control during Installation . . . 348
3.11.3 Checking Quality Control after Installation . . . . 357
3.12 Quality Control Improvement . . . . . . . . . . . . 359
Chapter 4
Verification: Cablebolt Performance Assessment
4.1 Introduction . . . . . . . . . . . . . . . . . 361
4.2 Visual Performance Assessment . . . . . . . . . . . 362
4.2.1 Remote "Visual" Data Collection . . . . . . . 363
4.3 Monitoring Performance with Instruments . . . . . . . . 366
4.3.1 The Instrument Toolbox . . . . . . . . . . 367
4.3.2 Design of the Instrumentation Program . . . . . 370
4.3.3 Installation of the Instruments . . . . . . . . 374
4.3.4 Data Recording . . . . . . . . . . . . 374
4.3.5 Data Reduction and Plotting . . . . . . . . 375
4.3.6 Data Visualization and Interpretation . . . . . . 377
4.4 Instrumentation and Failure Analysis . . . . . . . . . . 382
4.5 Experience: The Best Design Tool . . . . . . . . . . . 383
4.6 Performance Assessment Feedback for Cablebolt Design . . . . 384
References . . . . . . . . . . . . . . . . . . . . . 385
Index . . . . . . . . . . . . . . . . . . . . . . . 401
vii
Foreword
The need for this book arose during an industry sponsored research project on
rock support in underground hard rock mines carried out jointly at the Canadian
Universities of Toronto, Queen's and Laurentian between 1989 and 1993. The
topic of cablebolting could not be covered adequately within this broad research
field and so a follow-up research project was initiated by the Geomechanics
Research Centre, at Laurentian. This book is the end product of this research
project which involved visits to mines in Canada, Australia and Papua New
Guinea and a six month visit by the authors to the Rock Reinforcement Group of
the C.S.I.R.O. in Perth, Australia.
Jean Hutchinson and Mark Diederichs, who were graduate students and
research engineers in the Department of Civil Engineering at the University of
Toronto at the start of the first project, carried out this follow-up work as a wife
and husband team with the Geomechanics Research Centre in Sudbury. They
brought many new insights and their own brand of enthusiasm to the project. The
many hours of hard work which went into preparing this book are reflected in the
clarity of the text and tables and the excellent quality of the numerous illustrations.
Cablebolting in Underground Mines contains a wealth of information which
will be useful, not only to underground miners, but to anyone concerned with the
design of support for underground excavations for any purpose in any kind of
rock. This is the first book to bring together the practical details on grout mixes,
grout pumps, cable characteristics and the theoretical background required for the
rational design of cablebolt support systems. There are numerous mining and civil
engineering projects around the world in which cables are being used for support
and where the information contained in this book will be very valuable.
I commend the authors for their efforts in producing this fine volume and look
forward to using it in my own consulting and educational activities in rock
engineering.
Evert Hoek
Vancouver
October 1995

viii
Preface
Cablebolts are high capacity flexible tendons composed of multi-wire strand
which are normally installed and grouted in drilled holes at regular spacings to
provide reinforcement and support of excavations in rock. They can be cut to any
length and installed in single or multi-strand configurations and can be installed
from small adits and tunnels (drifts) where limited clearance would preclude the
use of rigid tendons.
The last decade has seen a dramatic increase in cablebolt usage in underground
mining. There has been a corresponding increase in associated research around the
world.
Researchers and industry experts came together for a day in June of 1992
during a cablebolting workshop held as part of the International Symposium on
Rock Support in Sudbury, Canada. Interest in the topic was impressive as
demonstrated by the standing-room only crowd at the workshop. In the aftermath
of this gathering, authors Mark Diederichs and Jean Hutchinson of the
Geomechanics Research Centre, together with Peter Kaiser and Dougal McCreath
of Laurentian University and with Chris Windsor and Alan Thompson of the Rock
Reinforcement Group (Commonwealth Scientific and Industrial Research
Organization, C.S.I.R.O., Australia), decided that there was a need to bring
together, in an easy-to-use and comprehensive handbook, the current state-of-the-
art in cablebolting.
Funding for the project, obtained from Canadian and Australian mining
companies, was coordinated by the Mining Research Directorate (M.R.D.) in
Canada and by the Australian Mineral Industries Research Association
(A.M.I.R.A.). The authors visited 50 mines in Canada, Australia and Indonesia,
observing mining techniques, ground conditions and cablebolting practice,
consulting with mine staff to determine the current state-of-the-art in cablebolting
and to tapping local expertise. The authors also communicated with numerous
international researchers in the field.
The result is this comprehensive handbook, covering virtually all aspects of
cablebolting, for support of underground excavations, from theory to practice with
an emphasis on (but not restricted to) applications in the mining industry. It is an
essential guide for the rock mechanics or ground control engineer in mining and
in civil construction and is an excellent reference for researchers and developers
in the field.
Recent innovations in cablebolting are presented and the current body of
international research is summarised in the context of cablebolt type selection and
support system design (Chapter 2). Installation and quality control procedures are
outlined along with suggestions for crew training and management (Chapter 3).
Practical techniques for support performance assessment and design verification
(Chapter 4) round out the "Cablebolting Cycle" as introduced in Chapter 1.
ix
Project Sponsors
Mining Research Directorate
Canada:
American Barrick Resources Grp.
Cambior Inc.
Cominco Ltd.
Falconbridge Ltd.
INCO Ltd.
Placer Dome Inc.
Williams Operating Corp.
Northern Ontario Development Agreement:
Canada Centre for Mineral
and Energy Technology,
Natural Resources Canada,
Ontario Ministry of Northern
Development and Mines.
Australian Mineral Industries
Research Association:
Ausdrill Pty. Ltd.
BHP Engineering
Billiton Australia
Dept. of Mines and Energy, N.T.
Hamersley Iron Pty. Ltd.
Murchison Zinc Co. Ltd.
Newcrest Australia Ltd.
Ok Tedi
Pancontinental Mining
Porgera Joint Venture
Rock Engineering Pty. Ltd.
Western Mining Corp.
Woodlawn Mines
Chief Advisors and Contributors
Geomechanics Research Centre and the Peter K. Kaiser
School of Engineering, Laurentian University Dougal McCreath
Rock Reinforcement Group, Commonwealth Alan Thompson *
Scientific and Industrial Research Organization Chris Windsor *
* currently with Rock Technology Pty. Ltd., Western Australia
Correspondence
The authors welcome from the reader comments on this book or suggestions
for future editions. These contributions and other correspondence can be sent to:
D. Jean Hutchinson & Mark Diederichs
c/o Geomechanics Research Centre
Fraser Building, F217, Laurentian University
Ramsey Lake Road, Sudbury, Ontario, Canada, P3E 2C6
email: mdiederi@nickel.laurentian.ca
For book ordering information please see BiTech address on page ii.
Disclaimer
The authors and their affiliated organizations, as well as the contributors, sponsors
and publisher of this handbook disclaim any responsibility for the applicability or
correctness of the information, recommendations and guidelines presented in this
document, or for the consequences resulting from the use thereof. Any use or misuse of
the information contained in this document is the sole responsibility of the user.

x
Acknowledgements
The contributions of the sponsoring companies and groups have made the
whole project possible, for which we are extremely grateful. Thanks to Charlie
Graham of the M.R.D. and to Jim May and Roger Wischusen of A.M.I.R.A. for
coordinating the necessary funding for this work.
We are deeply indebted to all of the very busy ground control and mining
engineers in Canada, Australia and Papua New Guinea, who guided us around
their mines, and who discussed their cablebolting design, installation procedures,
problems and solutions with us. This book could not have been written at all
without their input. The mine site visits conducted during this project provided us
with the necessary practical grounding for this handbook.
We are also pleased to have been able to work with a number of fine
researchers prior to and during the course of this project, all of whom provided
excellent opportunities for discussion and evolution of thought on cablebolts and
rock mechanics. During the course of the project we gained a great deal from
discussions with Will Bawden and Andy Hyett at Queens' University, Kingston,
Canada; John Goris of the U.S. Bureau of Mines, Spokane, USA; Phil Oliver,
formerly of INCO Mines Research, Sudbury, Canada; and Ernesto Villaescusa of
Mount Isa Mines, Queensland, Australia.
In Australia, the entire Rock Reinforcement Group provided truly five-star
hospitality and excellent technical assistance. Chris Windsor, Alan Thompson,
Glynn Cadby, Wayne Robertson, Pat Carden, Bob Middleton, Rosalie Thompson,
and Michele Currey made our stay in Perth a memorable personal and professional
venture. Thanks in particular to Chris and Alan for making it all possible.
At home at the Geomechanics Research Centre and at Laurentian University,
thanks are due to Peter Kaiser, Dougal McCreath, Dwayne Tannant and Véronique
Falmagne. Thanks also to all the staff at the G.R.C. for their support.
A number of people have assisted by reading and providing feedback on the
chapters of the book as they have evolved. At the risk of leaving someone out, we
would like to thank Jane Alcott, Gary Auld, Trevor Carter, Chris Langille, Peter
Mikula, Doug Milne, Marnie Pascoe and Norikazu Shimizu for their comments.
A very special acknowledgement is due to Evert Hoek who introduced us so
elegantly and brilliantly to the subject of rock mechanics and whose leadership
ultimately made much of this work possible.
Finally, we thank our families. The authors were married to each other just
before the start of this project and remain happily married at the end. While we
thank each other for the understanding and mutual support which made this
possible, our families were invaluable in keeping real life and work in perspective.
Figure 1.1.1: a) Cablebolt element b) Typical cablebolt array
1 INTRODUCTION:
Cablebolting in Underground Mines
1.1 What is a Cablebolt?
A conventional cablebolt is a flexible tendon consisting of a number of steel
wires, wound into strand, which is grouted into a borehole. Cablebolts are
normally installed in regularly spaced boreholes to provide reinforcement and
support for the walls, roof and floor of underground or surface openings.
Cablebolting is a very versatile form of support, since the cable strands can
bend around fairly tight radii, making installation of long bolts from confined
working places possible, and because they can be fabricated using a number of
different configurations of the steel wires providing a variety of performance
characteristics. It is not difficult to place more than one cablebolt strand in a
single borehole, to increase tensile capacity, if the borehole diameter is large
enough. In addition, face restraint can be attached in the form of plates, straps and
mesh. Cablebolts can be used in combination with other support systems such as
shotcrete, mechanical bolts or grouted rebar.
The capacity of the steel cablebolt element is transferred to the rockmass
through grout. Grout used in cablebolting applications is usually composed of
Portland cement and water. At some mine sites additives are added to the mix to
improve the pumping characteristics of the grout. Other grouts, including resin
and shotcrete have been investigated for certain cablebolting applications. As
well, alternative materials such as fibreglass have been developed to replace the
steel itself. This handbook will focus primarily on cablebolts developed from the
conventional seven-wire steel strand and on cement grouts.

2 Cablebolting in Underground Mines
1.2 Why Cablebolt?
Cablebolts are used in underground hard rock mines to:
+ provide a safe working environment,
+ increase rockmass stability, and
+ control dilution of waste rock from the stope boundaries.
In any mining or construction project, safety is of paramount importance.
Different support methods such as mechanical rockbolts and screen, shotcrete or
grouted rebar are normally employed in smaller span mining tunnels or drifts to
protect workers from smaller blocks and loose rock which may fall from the roof
or sidewall. For larger spans in major intersections, large underground chambers
or in active mining stopes, cablebolts become an attractive support system due to
the increased load capacity and the potential for increased bolt length. Larger
spans in general mean greater potential for large free blocks or broken rock falls.
This increased demand requires an increase in support system capacity which can
be effectively provided by cablebolts to ensure adequate safety.
It should be noted that in areas where areal restraint systems such as screen are
used to protect miners from smaller pieces of loose rock and to prevent surface
unravelling of poor rockmasses, cables can be used to supplement but not to
replace this form of restraint. This form of demand cannot be accommodated by
cablebolts alone. Chapter 2 discusses demand-capacity relationships.
Cablebolts can reach far into the rockmass and reinforce large volumes of rock
to prevent separation along planes of weakness such as joints. By maintaining a
continuum nature within the rockmass, the cablebolts help to mobilize the inherent
strength of the rockmass, thereby improving overall stability. In addition, by
supporting blocks of rock at the excavation surface, the remaining rockmass is
prevented from loosening and weakening. Cablebolts thus restrict the dangerous
and costly effects of progressive instability and failure.
Cablebolts can be installed remotely in long boreholes to reach the planned
stope boundary and provide pre-reinforcement to the otherwise inaccessible walls
and backs created by today's bulk mining methods. Cables are one of the only
options for support of inaccessible rock faces for stability and dilution control.
Dilution control can have a very direct and large influence on the cost of a
stope. The cost of dilution is many-fold: waste rock with little or no economic
value is mucked, trammed, crushed, skipped, milled and impounded in a tailings
disposal area, all at great cost. In addition, the mill works at effectively only
partial capacity, despite producing at the maximum possible milling rate. The
unscheduled delays required to deal with oversize muck, mucking waste rock and
with consequent changes to the mining schedule are also costly.
Introduction 3
Figure 1.2.1: Definition of terms
(after Anderson and Grebenc, 1995)
Figure 1.2.2: Cablebolting pattern and dilution surveyed in adjacent stopes at Hemlo
Gold Mine (after Anderson and Grebenc, 1995)
Anderson and Grebenc (1995) provide an excellent discussion of the several
components of dilution as shown in Figure 1.2.1 and defined here. The required
information is collected from a laser survey of each stope after mining is complete
(see Section 4.2 for further discussion of laser stope surveying). Factors which
must be considered in assessing the performance of a stope include:
% Dilution =
Waste dilution (t) + Backfill dilution (t)
x 100
Planned tonnes (t)
% Recovery =
Planned tonnes (t) - Ore lost in stope (t)
x 100
Planned tonnes (t)
% Overbreak =
Ore sloughing (t)
x 100
Planned tonnes (t)
Dilution control has a high priority at
Hemlo Gold mines. Every stope is surveyed,
so that the factors listed above can be
calculated. Anderson and Grebenc provide a
very illustrative case history of dilution
control through the understanding of the cause
of failure in one stope and the effective design
of support (cablebolt and drift backfill) for the
adjacent stope. The difference in the
cablebolting pattern and dilution of the stope
walls is shown in Figure 1.2.2.

Figure 1.2.3: Dilution vs sloughing & span
(after Pakalnis et al., 1995)
Figure 1.2.4: Economic impact of dilution
(after Bawden, 1993)
Other examples illustrating the economic importance of rock dilution are given
throughout the mining literature including Bawden (1993), Elbrond (1994),
Pakalnis et al. (1995), Planeta et al. (1990), Planeta and Szymanski (1995),
Stillborg (1986), and others. Many mining handbooks include calculations for
tracking the progress of waste rock dilution through the mining and milling
processes to determine its overall economic impact. A detailed treatment of mine
economics, however, is beyond the scope of this handbook.
All modern mining will have some minimal dilution limit resulting from the
smoothing of stope outlines to facilitate blasting or due to other sources of planned
dilution. In many situations, particularly where there is a distinct ore/waste
contact, unplanned dilution due to sloughing waste rock can quickly render the
stope uneconomic. It is this unplanned dilution component which can be tackled
through improved stope design and through the use of cablebolt support.
Many mines have found
cablebolts to be effective in
reducing or eliminating this
sloughing and thereby reducing
dilution. Alternatively, the use of
cablebolts can facilitate the safe
extraction of larger stopes
normally resulting in increased
productivity. Figure 1.2.3 shows
theoretical dilution values as a
function of span and sloughing
depth for an unsupported stope of
simplified geometry.
As this dilution moves through
the system incurring additional
mucking, haulage and hoisting
costs as well as (and most
importantly) displacing profitable
ore (grade reduction) in the mill
(Bawden et al., 1989), it becomes
apparent that the economic
consequences can be extreme
(Figure 1.2.4). Additional costs and
losses are incurred due to the effect
of unplanned downtime required to
handle oversized waste rock. In
areas where cablebolting is effective
in reducing dilution, the cost of
cablebolting (Section 1.9) is often
minuscule by comparison.
Introduction 5
Figure 1.3.1: Cablebolt layout examples for mining excavations
1.3 Cablebolt Applications
Cablebolts can be used to support, reinforce or retain the rockmass around
most excavations found at an underground mine, including:
+ Drifts and intersections.
+ Open stope backs.
+ Open stope walls.
+ Cut and fill stopes.
+ Drawpoints.
+ Permanent openings.
The particular cablebolt pattern selected will depend upon the intended function
of the cablebolts and the access for installation. Access for cablebolting is usually
provided by production drifts. While the cost of driving drifts solely for installing
support is expensive, a number of mine sites have reduced the extent of rockmass
failure around open stopes (and thereby dilution) by installing more effective
cablebolt patterns from "cabling drifts" (see Figure 1.2.2). Some sketches showing
examples of cablebolt layouts for different stope and access configurations are
shown below. These patterns can be used individually or in combination to
provide the most effective cablebolt pattern.

Figure 1.3.2: Example cablebolt applications and layouts
Cablebolt Applications
Introduction 7
Figure 1.3.3: Example cablebolt applications and layouts
Cablebolt Applications

Figure 1.4.1: The cablebolting cycle
1.4 The Cablebolting Cycle
The Cablebolting Cycle shown in Figure 1.4.1 provides a comprehensive
overview of the steps involved in the cablebolting operation. It also forms the
basis of the structure of this handbook: Chapter 2 discusses Design, Chapter 3
outlines Implementation and Chapter 4 covers Verification.
The cycle displayed on this diagram represents a cyclical, iterative process
which should be worked through a number of times as mining progresses to ensure
that the cablebolting process is well-tuned.
Introduction 9
1.4.1 Design
The design of any system is based on efficiently matching the available
capacity with the required demand while remaining bounded by certain specified
constraints. In the case of excavation support design the demand arises from the
disturbed rock mass, after the equilibrium of stress and of gravity is disturbed by
the creation of an opening. The rock must move to adjust to a new equilibrium. If
this is not possible, support must aid in achieving this equilibrium. This requires
support properties such as stiffness, load capacity and load displacement capacity.
Cablebolt capacity is discussed at length in Chapter 2. Key aspects to consider:
+ Loading configuration and testing (Sections 2.2 and 2.8)
+ Cable array (Section 2.3)
+ Strand capacity (Section 2.4)
+ Grout strength / stiffness (Section 2.5)
+ Cable/grout and grout/rock bond capacity (Section 2.6)
+ Plating and surface fixtures (Section 2.7)
Demand assessment is covered in Chapter 2 using
+ Empirical approaches based on experience (Sections 2.14 to 2.17)
+ Mechanistic approaches based on behavioral analysis (Section 2.18)
Constraints are placed on the support system design by the economics of
mining including cablebolting costs and potential losses due to instability. Other
constraints include regulated safety standards, mining and development sequences
and access for equipment. The determination of a cablebolting configuration
including equipment selection is covered in Sections 2.9 to 2.12.
1.4.2 Implementation
Implementation of cablebolts, as discussed in Chapter 3, involves effective
communication and quality control in addition to installation. Communication is
critical to the successful implementation of a support system. Recommendations
for crew instruction, training and communication are discussed in Sections 3.2 to
3.5. Installation procedures are detailed in Sections 3.6 to 3.10. Quality control
guidelines and monitoring procedures are presented in Section 3.11 and 3.12.
1.4.3 Verification
Chapter 4 briefly discusses verification and performance assessment. In order
to justify the expense of a support system and to optimize the efficiency of a
particular design, a verification program must be implemented, consisting of
visual observation (Section 4.2) where possible, instrumentation (Section 4.3) and
measurement of rockmass and support performance and must include feedback
(Section 4.4 to 4.6) into the design process.

1.5 The Cablebolt Toolbox
The cablebolt toolbox includes a wide variety of items that allow the user to
design a truly effective cablebolt element for most potential rockmass failure
conditions. The toolbox includes a number of different cablebolt strand
configurations and different grouting materials. Additional items in the cablebolt
toolbox include surface restraint elements such as plates and straps.
The basic cablebolt that has been used around the world for a number of years
is the plain strand cablebolt. In the last 20 years, a number of different types of
modified cablebolt strand have been developed in response to problems
encountered with poor performance of plain strand cablebolts at mine sites. Some
of these modified geometry cablebolt strands are shown in Table 1.5.1. Further
discussion of the characteristics of the modified strands is made in Section 2.9.
Cablebolts have traditionally been grouted with a grout composed of cement
and water. The grout water:cement ratio observed at mine sites by the authors
ranges from 0.3 to 0.6 or 0.7. Laboratory cablebolt pull tests on plain strand
conducted at the US Bureau of Mines by Goris (1990) and at Queen's University
by Reichert, Bawden and Hyett (1992) have shown that the water:cement ratio
(W:C) of the grout should be kept in the range of 0.3 to 0.4 for optimum cablebolt
capacity and performance. The lower W:C grout will give the best capacity in pull
out tests. However increasing scatter of laboratory UCS test results at W:C = 0.30,
problems with mixing, pumping and grouting very thick grouts with some
equipment, and concerns about the ability of very thick grouts to flow into the
cages of modified strands, create conditions in which a thinner but adequately
strong grout (W:C = 0.35 - 0.4) is specified in design.
A detailed discussion of the influence of grout strength on cablebolt capacity
is provided in Sections 2.5 and 2.6. A schematic summary of grout mix design is
given in Figure 2.5.9.
Other materials and concrete admixtures have been used to grout cablebolts at
some mines. Epoxy based grouts have been used in highly corrosive environments
to protect the steel strand. Sanded grouts and shotcrete grouts have also been
investigated in laboratory tests.
Plates and straps are used to "tie" the cablebolt element to the exposed
rockmass surface. This is important in areas where the surface rockmass is not
sufficiently retained by the linear cablebolt element, and where there is access to
the "working end" of the cablebolts. In this case, the cablebolts must be correctly
tensioned during installation of the surface retaining elements which must be
securely attached to the end of the cablebolt with a wedge and barrel device
(Section 2.7).
Introduction 11
Table 1.5.1: The cablebolt toolbox (after Windsor, 1992). Detail in Section 2.9.
Longitudinal Section Cross Section
Single
plain strand
Double
plain strand
with spacers
Birdcaged
strand
Bulbed
strand
Ferruled
strand
Nutcaged
strand
Epoxy-
coated or
encapsulated
strand
Buttoned or
swaged
strand

Figure 1.6.1: Primary support functions
(after Kaiser et al., 1995)
1.6 Cablebolt Function
Cablebolt support performs a combination of reinforcement and holding
functions. As reinforcement, the cables prevent separation and slip along planes
of weakness in the rockmass. If moderately rough joint and fracture surfaces can
be kept from separating, the influence of these discontinuities can be minimized.
An effectively continuous rockmass is almost always stronger than a discontinuous
one and therefore cablebolts help to mobilize the inherent strength of the jointed
or fractured rockmass. Cablebolts cannot, however, increase the overall rupture
strength of a continuous rockmass and are unlikely to prevent hard rock from
fracturing under high stress. If the inherent strength of the rockmass is not enough
to resist the effect of induced stresses or if discontinuities are unfavourably
oriented resulting in free and removable blocks, cablebolts can be effective
holding elements, keeping the failed rock or free rock blocks in place.
Cablebolts are inefficient retention
(maintenance of small loose surface
particles) elements in poor quality
rockmasses unless used in combination
with screen, shotcrete, straps or other
surface coverage. The same is true for
fractured rock at depth. If the fracturing is
intense, the cablebolts may not be able to
retain the fractured rock in place. If the
rockmass at the excavation surface is held
together by other surface retention systems,
cables attached to the retainer elements can
provide effective holding capacity.
Hard Rock vs Soft Rock
This handbook was developed to suit the needs of the hard rock mining
industry. According to Carter (1995), hard rock includes most igneous rocks, non-
schistose metamorphic rocks and well cemented sedimentary rocks. Soft rock
includes highly weathered or altered igneous rocks, weakly cemented rocks such
as schists, phyllites, shales, silts and fault zones. Although coal is often considered
to be soft rock, most of the guidelines and recommendations in this book apply to
many coal mining environments. Some so-called hard rock mining environments
can contain zones of very poor quality (talc schists, altered dykes). In general, the
guidelines and recommendations contained in this handbook can be applied, with
suitable caution, to most rock types encountered in underground mining. An
exception may be viscous materials such as salt and potash. Support logic
described in this book can only be applied to the short term behaviour of these
materials. Excavations with excessive water inflow in poor quality rock and with
extremely poor, squeezing soil-rock are also not covered here.
Introduction 13
Figure 1.6.2: Typical cablebolt functions (italics indicates cablebolt function only)
Cablebolt Function

1.7 Equipment
There are a number of different cablebolt installation operations that require
specialized materials and equipment. This equipment might have been purchased
for other applications at the mine site, but generally is uniquely useful for
cablebolting applications. Wherever possible, it is advisable to dedicate specific
equipment units for cablebolting and to make the maintenance and care of these
units the responsibility of the cablebolt crew. Regular maintenance and post-shift
clean-up can be made an integral part of the cablebolting payment incentive
system to ensure minimum unscheduled downtime.
The equipment is discussed in a number of places in this handbook, so will be
summarized in a list here for reference. Some of the items in the list will be
required at all sites, and some are designed for a specialized task that might be
carried out at all sites. This is not an exhaustive list, and there are likely to be
alternative types of equipment that are better for certain operations.
In addition to the cablebolt materials including strand (Sections 2.4 and 2.9),
grout (Section 2.5), plates and anchors (Section 2.7), tubing (Section 2.10 and
2.11.3) and attachments (Section 2.11.4), the cablebolting equipment list could
include:
+ Stationary cablebolt reel or revolving dispenser for dispensing the cablebolt
(Section 2.11.4).
+ Hydraulic cutter, air powered grinder, oxy-acetylene torch, or explosives for
cutting the cablebolt (Section 2.11.4).
+ Custom built cablebolt pushers (Section 2.11.4).
+ Paddle, drum or colloidal grout mixers (Section 2.11.2).
+ Piston or progressing cavity grout pumps (Section 2.11.2).
+ Tension jack for tensioning the cablebolt during surface element (plate, strap)
installation (Section 2.7).
+ Cablebolting truck equipped with all of the items listed above, and any tools
required by the crew (Section 2.11.4).
In the experience of the authors, quality control and productivity are both
greatly enhanced by a well equipped, self-contained, mobile and organized
cablebolting crew. In general, where cablebolting is to be a major ground control
priority, time and money spent creating an efficient, well trained cablebolting unit
will pay dividends.
Introduction 15
1.8 Cablebolt Installation Options
There are a number of different methods in use for grouting cablebolts. The
selection of the best method depends upon the orientation of the borehole, the type
of cablebolt, the grout flow characteristics and the grouting equipment available.
A brief description of the different installation methods is given in the
following pages. Installation methods are discussed in more detail in Section 2.10
and in Chapter 3. The cablebolt installation methods most commonly used are:
+ Breather Tube Method. This method is used for upholes only and with grout
of 0.375 - 0.45 water:cement ratio. The optimum grout for this installation
method is 0.4 W:C. (The consideration of other important items such as
cablebolt types, breather tube diameter etc. may alter the range of optimum
grout water:cement ratio specified in design). This method should be used with
caution in areas with open fractures in the back which may cause grout loss
with thinner grouts and may prevent complete filling of the hole. In loose,
thinly laminated ground, caution is required to avoid over pressurizing fractures
causing the laminations to separate and rupture.
+ Grout Tube Method. This method can be used for any hole orientation and
with grout of 0.30 to 0.375 water:cement ratio. The optimum grout for this
installation method is 0.35. These thicker grouts may cause pumping
difficulties with less powerful pumps and long holes. The use of modified
geometry cablebolt elements may require slightly wetter grout at W:C of 0.37.
In both of these methods, the tube(s) are attached to the cablebolt strand prior to
the placement of the cablebolt in the borehole. The grout front flows along the
entire length of the borehole in these methods.
Variations on the grout tube method in use at some mines are:
+ Retracted Grout Tube Method. The grout tube and cablebolt are placed in the
borehole and then the reusable grout tube is withdrawn from the borehole as the
grout is being pumped. The grout is placed in position in the borehole and does
not flow over appreciable distances. Care and experience is required to prevent
void formation in the grout column. A highly skilled operator is required.
+ Grout and Insert Method. In this method the borehole is grouted using the
Retracted grout tube method and then the cablebolt is pushed into the grout
filled borehole. The diameter of the borehole can be reduced in this method,
and the grout tube is reusable. This method is generally used with automated
cablebolting equipment. Overly rapid tube withdrawal or cable insertion will
result in a poorly coupled system.
Each of these installation methods is described in the following pages. The
method most likely to completely grout the hole should be selected.

1.8.1 The Breather Tube Installation Method
+ In this method, the breather tube extends to the toe of the hole, while only a
short length of grout tube is used at the collar of the hole. A cablebolt hanger
and borehole collar plug are required.
+ Grout of 0.4 water:cement ratio is optimum for this method.
+ The grout is pumped through the short grout tube into the borehole. The grout
flows upward against gravity in the hole. Air and then grout are expelled from
the hole through the breather tube. Return of good quality grout through the
breather tube is essential to indicate that the borehole is full of grout.
+ A piston pump or progressing cavity pump can be used.
+ Problems encountered with this method include: leaking or blown out collar
plugs, caused by poorly plugged collars or undersized breather tubes, grout
much wetter than design consistency; and no grout flow from the breather tube
due to loss of grout into a badly fractured rockmass, an undersized breather
tube for the design grout consistency, or inadequate pumping time.
Introduction 17
1.8.2 The Grout Tube Installation Method
+ The grout tube extends to the toe of the hole. A cablebolt hanger at the toe
and/or a wooden wedge inserted at the collar secure the bolt in upholes.
+ Grout of 0.37 water:cement ratio should be used for upholes.
+ In upholes, the grout is pumped to the toe of the hole through the grout tube.
The grout then flows downward with gravity inside the borehole. The grout
must be thick enough so that at the instant the pump is stopped, the position of
the grout flow front will freeze in the hole. A thick consistency donut of
grout appearing at the collar indicates complete grouting of the hole.
Obstructions, such as the wires of a modified cablebolt strand or spacers, may
divide the grout front, leaving voids in the grout column.
+ A continuous stream of grout is required, so a progressing cavity pump is
usually used.
+ Voids can easily be created in upholes: too thin grout will slump or spiral down
the hole, and thick grout may hang up in the hole preventing complete grouting.

1.8.3 The Retracted Grout Tube Installation Method
+ The grout tube extends to the toe of the hole, but is retracted and can be reused.
A cablebolt hanger is required to secure the cablebolt in upholes.
+ Grout of 0.37 water:cement ratio should be used for upholes.
+ The grout is pumped to the end of the grout tube, which is withdrawn slowly
from the borehole. In this method, the grout is placed at the required position
and flows only a short distance within the borehole. If the grout tube is
withdrawn too quickly, voids will be created in the grout column. The grout
must be thick enough so that it will hang up in an uphole. This method is the
most reliant of the four on good crew skills and training.
+ The pump must have enough power to pump thick grout into the longest hole.
+ Voids are easily created: too thin grout will slump down upholes, and too thick
grout may freeze in the grout tube. If the grout tube is withdrawn from the hole
too quickly, voids will also be left in the grout column.
Introduction 19
1.8.4 The Grout and Insert Installation Method
This method is generally used for cablebolting machines only, since a lot of
force is required to push a cablebolt through the column of grout.
+ In this method the reusable grout tube is pushed to the end of the hole, then is
retracted during grouting. The cablebolt is inserted into the grout filled hole.
+ Grout of 0.37 to 0.35 water:cement ratio should be used for upholes.
+ The grout is pumped to the end of the grout tube, which is withdrawn slowly
from the borehole. In this method, the grout is tremmied into place so that it
flows only a short distance within the borehole. The grout must be thick
enough so that it will not slump down in upholes (W:C 0.37), but not so thick
that it will not fully encapsulate the cablebolt strand.
+ The pump must have enough power to pump thick grout into the longest hole.
Either a piston or progressing cavity pump can be used.

1.9 The Cost of Cablebolting
The cost of cablebolting varies and can often seem high, relative to other types
of support. However, if the cablebolts have been well designed and installed, they
should reduce mining costs appreciably by reducing expensive dilution costs, more
than paying for themselves. In addition, cablebolts that are performing well will
increase the safety of the people working in the mine, and will increase the
stability of the immediate and of the surrounding mining excavations. Some costs
for cablebolting are included here for general information.
Table 1.9.1: Unit cost for cablebolting including drilling (after Goris et al., 1994)
Mine Cablebolt applications at different mine
sites
Cost (1992 $ Canadian)
$ / m of cablebolt
A Single cablebolts with 0.3 by 0.3 m plates $29.46
B Double cablebolts 28.84
C Double cablebolts 28.54
D Single cablebolts 31.83
E Single cablebolts 19.69
Table 1.9.2: Typical cost for a 12.2 m long twin strand cablebolt (after Goris et al., 1994)
Item Cost (1992 $ Canadian)
$ / cablebolt
Hole drilling, including labour 145.67
Twin cablebolt strand 44.04
Cablebolt hanger 4.25
0.3 by 0.3 m steel plate 2.45
Wedge and barrel 3.30
Grout tube to toe of hole 12.46
Cement 10.00
Labour 52.09
Total cost 274.26
Average cost per metre 22.48
Table 1.9.3: Production rates for cable insertion and for grouting (after Goris et al., 1994)
Mine Crew Size Cable
Length
m
Grout
Water:Cement
W:C
Productivity rate
meter of cablebolt / shift
Cable Insertion Grouting
F 1
2 9 - 15 0.32:1 90 180
G 1
2 15 0.45:1 69 166
H 2
3 15 - 20 0.32:1 166 230
NOTE: 1 - Mines F G; 8 hour shift. 2 - Mine H; 12 hour shift (Drilling not included)
Introduction 21
1.10 A Note About Units
This handbook uses S.I. units exclusively. This is to avoid confusion and to
simplify computation. The following conversions are provided for convenience:
Distance (Length, Width, etc.)
1 foot (ft) = 0.3048 metres (m)
1 inch (in) = 0.0254 metres (m) = 25.4 millimetres (mm)
1 metre (m) = 1000 mm = 3.2808 feet (ft) = 39.3701 inches (in)
Area
1 square foot (ft or sq. ft.)2
= 0.0929 square metres (m or sq. m.)2
1 square inch (in or sq. in.)2
= 0.000645 (m ) = 645.16 (mm )2 2
1 square metre (m or sq. m.)2
= 10.7639 (ft )2
Volume
1 cubic foot (ft or cu. ft.)3
= 0.0283 cubic metres (m or cu. m.) = 28.316 litres (l)3
1 cubic inch (in or cu. in.)3
= 1.6387x10 m = 0.0164 l-5 3
1 litre (l) = 0.001 m3
= 0.0353 ft = 0.02642 U.S. Gallons3
= 0.21998 U.K. Gallons
Mass
1 kilogram (kg) = 2.2046 pounds (lbs) (mass)
1 tonne (t) = (1000 kg) = 2204.622 pounds (lbs) (mass) = 1.1023 tons (short)
1 short ton = 2000 lbs (mass) = 0.9072 tonnes (t)
Force
1 Newton (N) = 0.2248 pounds (force)
1 kN = 1000 kg.m.s-2
= 224.91 lbs (mass) x 1 g (gravitational acceleration)
Pressure
1 MegaPascal (MPa) = 145.05 pounds (force) per square inch (PSI)
1 PSI = 6.895 kN/m = 6.895 kPa2
Note: Force - Mass equivalency:
Under gravitational acceleration (1g = 9.81 m/s ), 1000 kg of mass (1 tonne)2
produces 9.81 kN of force. In this manual, the relationship, 10 kN = 1 tonne, is
used for simplicity and familiarity. The 2% error is not significant, but it is
important to remember that tonnes represent mass, while kiloNewtons represent
force (mass x acceleration), and that 10 kiloNewtons is the approximate force
generated by 1 tonne of mass at 1 g of acceleration (due to gravity). Also
remember that a metric tonne is approximately 6% larger than an Imperial ton.

1.11 Useful Definitions
Density, Unit Weight and Specific Gravity
In engineering applications, these terms are often incorrectly used
interchangeably to describe the relative heaviness of materials. It is important to
understand the differences in the meaning of these terms in order to avoid critical
errors in design calculations.
Density, ,,
Density describes the amount of material or mass that is contained within a
specific volume. One litre of water, for example, contains one kilogram of the
liquid. One cubic metre contains one thousand kilograms or one metric tonne of
liquid water. Therefore we say that the density of water is one tonne/m or one3
kilogram/litre. A felsic granite, for example has a density of 2.7 tonnes/m , while3
a high grade sulphide can have a density greater than 3.3 tonnes/m .3
Unit weight,
Unit weight describes the weight or force exerted by gravity on a unit volume
of material. It is obtained by multiplying the density by gravitational acceleration
or 9.81 m/s . (A more convenient conversion factor of 10m/s can be used for2 2
most practical applications). The resultant value is expressed most conveniently
in units of kiloNewtons/m (kN/m ) or MegaNewtons/m (MN/m ). The unit3 3 3 3
weight of water, therefore, is given by:
1000 kg/m x 9.81 m/s = 9810 N/m = 0.0098 MN/m3 2 3 3
Granite has an approximate unit weight of 0.027 MN/m , and a high grade3
sulphide an approximate unit weight of 0.033 MN/m .3
Specific Gravity, S.G.
The specific gravity of a material is simply the dimensionless ratio of either the
unit weight or the density of a material to the respective unit weight or density of
water. The specific gravity of water is, of course, unity or 1. The specific gravity
of granite becomes approximately 2.7, while the sulphide has a specific gravity of
approximately 3.3. Note the absence of units. This is therefore a convenient term
to state the relative heaviness of materials, since it is independent of the system of
measurement and the units used.
When performing calculations in this book for hard rock underground
applications, it is convenient to use a specific gravity of 3.0 or a unit weight of
0.03 MN/m , if the true value is not known. This value is an average between3
barren waste rock and higher grade metallic ore. Figure 2.1.1: Key considerations for cablebolt design in underground mines
2 DESIGN
Application of Engineering Principles
2.1 Introduction
This chapter summarizes most of the key considerations involved in the design
of cablebolt systems in underground mining environments.

2.1.1 Design Acceptability Criteria
Safety
In areas with high occupational or flow-through traffic such as refuge and shaft
stations, garages, crusher and equipment rooms and haulageways, safety is likely
to be of paramount importance. While small pieces of loose will go unnoticed in
an open stope, they pose a serious hazard where personnel are present. In using
this handbook, supplement designs as needed to ensure adequate safety.
Stability
Under the influence of stress, gravity and vibration, rock can strain, yield,
deteriorate and ultimately disintegrate. Instability and failure can be defined as any
limiting point in this progression. Permanent critical openings such as crusher
stations and shafts may require a no-damage (yield) criterion while in temporary
drifts, time dependent collapse may be acceptable. Stability concerns can be local
(serviceability, access, rehabilitation costs) or can be global (destabilization of
mining block, pillar collapse, shaft pillar integrity, subsidence, etc.). The
consequences of instability should be evaluated as a part of mining engineering.
Dilution
In open stopes, it is not economically practical to attempt to prevent all forms
of instability. Limited dilution (waste rock overbreak or minor sloughing) is often
accepted within economic limits. The costs and revenue loss due to such
predictable dilution must be weighed against the costs of support (materials,
delays, labour) in order to determine the course of action - support or no support.
In non-entry open stope design, the decision whether (or not) to support at all can
have the most financial impact in the support design process.
Stand-up Time
Rockmasses are subject to time dependent deterioration in the vicinity of
excavations. An opening may be initially stable but may degrade over time,
eventually becoming unserviceable. The required stand-up times (supported and
unsupported) of an excavation should be established and considered in design.
Factor of Safety
Design calculations (bolt spacing, length, critical span, stand-up time) are often
based on numerous assumptions. In addition, the uncertainty inherent in the
measurement or collection of engineering data as well as the variability of the
underground environment mandate the use of a safety factor (multiplier, additive,
percentage,etc.) as appropriate to ensure that safe margins are built into the design.
Design: Application of Engineering Principles 25
2.2 Capacity and Demand
2.2.1 Introduction
In this handbook, the term capacity is used in a very general fashion. It is used
to encompass all aspects of cablebolt performance. In this chapter, the
performance specifications and expectations for many of the more common and
available cablebolt configurations are presented. Particular attention will be paid
to the plain strand (seven-wire) cable, since all other modified geometry strands
were developed to overcome deficiencies in the plain strand cablebolt.
Demand considerations address the necessary enhancements required to
stabilize an excavated rockmass. These arise as a result of excavation size and
geometry, the strength and structural integrity of the rockmass, the induced
stresses around the excavation and the aggressive and/or changing nature of the
excavation environment. The issues of demand are addressed through rockmass
classification, empirical and mechanistic design, local experience and through
rockmass monitoring. The demand requirements must then be matched to the
capacity of the selected support system.
Support systems based on the seven-wire plain or modified strand cablebolt are
primarily frictional, fully coupled devices. That is, load is transferred to/from the
rockmass along their entire length unless debonded sections occur as a result of
design or installation error. This transfer occurs as a direct result of friction
between the cable strand and the encapsulating grout. Load must also be
transferred between the grout and the surrounding rock. Modified strand has been
developed to increase the degree of cable-grout interlock, thereby increasing the
efficiency of the load transfer. If this so-called bond is optimized, then the
performance of the cablebolt system will be controlled by the quantity of steel
strand (cablebolt distribution), the geometry of the cablebolt array and
environmental changes after installation. The overall performance of the cablebolt
system can be subdivided into five capacity categories which can be directly
related to the categories of demand which they address. These are listed on the
following pages.
In most cases the performance data presented is the result of a synthesis of
available testing results from the literature and from unpublished contributions.
Every attempt has been made to simplify data into a practically useable form for
design and system selection. While the information contained here should be
sufficiently self-contained for preliminary design, references are cited for those
who wish to delve deeper into the body of literature regarding component
properties, behaviour, performance and testing. Cablebolt research continues to
be a healthy international industry and it is profitable to keep up to date with new
developments through industry and research publications and through suppliers.

Figure 2.2.1: Capacity considerations
Capacity Considerations
The capacity of a cablebolt element is based on the properties of the strands, on
the bond and frictional resistance of the interface between the cable and the grout,
on the quality of the grout and on the load transfer between the cable and the
surrounding rock. The capacity of the cablebolt support system is also determined
by the cablebolt pattern density, orientation and length. The system capacity can
be expressed as the sum of the following considerations:
+ Immediate stiffness describes the
relationship between initial loading
increments and their associated
displacements (cable stretch + interface
slip) within the cablebolt system. Of
significance is the stiffness over the first
1 - 10 mm of displacement.
+ Ultimate ductility describes the
maximum displacement that can be
accommodated by the cablebolt system
before total bond failure or cable strand
rupture. High ductility or displacement
capacity is desirable in highly stressed
or dynamically active ground.
+ Ultimate load capacity describes the
maximum static load which can be
sustained by the cablebolt before strand
rupture or total bond failure (free slip).
+ Surface retention is required to ensure
the local integrity of a rock face and/or
to guarantee personal safety. Cablebolts
spaced too far apart may permit face
disintegration between the cables,
requiring additional surface fixtures.
+ Longevity and sensitivity are important
considerations when cablebolts are
exposed to corrosive environments,
blasting and changes in local stress and
confinement. In addition, creep or
relaxation within the cablebolt system
can impair the support effectiveness.
Figure 2.2.2: Demand considerations
Demand Considerations
Demand assessment is based on a systematic consideration of the key rockmass
properties (initial stress, strength, stiffness, structural integrity, etc.) and of the
expected disturbing influences (stress change and gravity loading). Potential
failure modes can then be identified and where necessary, support can be designed
to maintain stability. Demand can be separated into five main components, each
corresponding to the respective capacity component on the previous page:
+ Dilation control; rockmasses normally
contain joints and fractures. These
surfaces are primarily frictional (shear
strength dependent on normal pressure)
and dilational (open during shear due to
roughness). If these surfaces can be held
together by stiff reinforcement, the
interlocking roughness and frictional
strength of the rockmass are maintained.
+ Displacement; where high stresses exist
or where smooth and continuous
discontinuities allow for large amounts
of slip under stress, displacements in the
rockmass may be severe. These stresses
are normally shortlived, however, as
stresses redistribute, leaving free
deadload blocks or broken rock zones.
+ Gravity loading of the support system
will dominate design if the rockmass is
allowed to disintegrate or if structural
features form free deadload blocks.
+ Surface ravelling can occur when
cablebolts are spaced too far apart with
respect to the block size in the rockmass
or if the near-face cable segments
possess inadequate bond strength.
+ Service life and robustness; corrosive
mine water can impair long-term
capacity of cablebolts. Stress and
confinement reductions in the rockmass
can impair rockmass stability and also
can reduce cablebolt capacity.

Figure 2.2.3: In situ loading of cablebolts (after Windsor Thompson, 1993)
2.2.2 Loading Configurations for Cablebolts
The first three categories of capacity and demand can also be subdivided into
subcategories of loading type:
+ Axial or tensile
+ Shear
+ Combination axial/shear
These modes of cablebolt loading occur individually or in combination within
an array of cablebolts as illustrated below. It is important to estimate the most
likely direction of motion (not always down or along a joint) in order to identify
the operative loading modes.
In Situ Loading of Cablebolts
Testing Configurations for Cablebolts
Axial loading tests are relatively simple to design and perform in the laboratory
and are also possible in the field. For this reason, the vast majority of available
data on cablebolt performance deals with axial tension testing.
Unconstrained tests (Maloney et al.,1992) are easily carried out but because
they allow the cable to rotate, they tend to give a lower bound strength. In these
tests a length of cable is grouted into a rigid pipe (see Hyett et al., 1992 for
equivalent stiffness relationships for pipe sections) with one long free length, for
gripping and pulling, at one end and a short free length at the other. To ensure a
constant embedment length, this shorter length should be equal to or greater than
the desired pull-out displacement. The long free end is used for gripping in the test
machine. Use grips which will not cause premature rupturing of the cable.
Displacements should be measured between a marked point on the cable and the
exit end of the grouted cylinder as shown in Figure 2.2.4.
Non-rotating (Goris, 1990; Hyett et al., 1992) and double pipe (Windsor and
Thompson, 1993; Villaescusa et al., 1992) tests are slightly more complex and
tend to give an upper bound on pull-out strength. The procedures for these tests
are detailed in these references. Note that the cable does not rotate during pullout
as it does in the unconstrained test. This forces the cable to shear through the grout
flutes and increases pull-out resistance. Displacement should be measured between
the opposing exit points in the two pipe sections as shown in Figure 2.2.4.
The difference between the constrained non-rotating single pipe test (Hyett et
al., 1992) and the double pipe tests is that in the former, the fixed section of pipe
is considerably longer than the test section and/or a swaged or welded anchor is
placed on the cable within the grouted test section to prevent slip. In the double
pipe test, both sample or test sections are designed to slip equally. As such, the
measured displacements in the double pipe tests will be twice those in the fixed
pipe constrained test as shown in the example result in Figure 2.2.4. It is important
to be aware of this effect when comparing results from different tests.
When performing these tests for the purpose of comparison between different
grouts and cable configurations it is important to record the following information
and to maintain control over those parameters which are to remain constant:
+ Cable Type
+ Grout W:C, Curing time and if possible UCS (samples from same grout batch)
+ Embedment length; Test type; Anchor length or free pull length
+ Pipe material and dimensions (Borehole diameter and properties for field tests)
+ Approximate pull rate
+ Pullout load and the displacement during the test.
+ Cable response notes (e.g. rotation, stick slip, cable strand rupture)

Figure 2.2.4: Three basic configurations for axial pull-out tests (laboratory)
Testing Configurations for Cablebolts - Axial
Figure 2.2.5: Axial field tests (after Maloney et al., 1992; Bawden et al., 1992)
In situ field tests can be carried out as shown in Figure 2.2.5. Unconstrained
tests are documented in Maloney et al. (1992). Note that the downhole length is
covered by a plastic tube, for debonding, except for the test section (test
embedment length). The entire hole can be grouted if desired. Constrained field
tests are more complex due to difficulties in constraining the down-hole cable
length. A procedure for constrained tests is detailed in Bawden et al. (1992). Note
that in either case, displacements will include cable (or loading rod) stretch.

Figure 2.2.6: Direct shear and combination (shear + axial) testing for cablebolts
(after Windsor and Thompson, 1993)
Testing Configurations for Cablebolts - Shear
Direct shear (Windsor and Thompson, 1993) and combined axial and shear
tests (Hyett et al., 1995) are complex and require specialized laboratory
equipment. It is also a complex procedure to properly simulate the shearing and
borehole conditions necessary for accurate results. Useful comparisons can be
made, however, between different cable systems and between the performance
(stiffness and load capacity) of cable strand with respect to loading angle. The
actual performance of the cable is dependent on the sense of the displacement
(shear, dilation and combined) and on the orientation of the cablebolt with respect
to the test interface and the direction of motion.
Figure 2.2.7: Combined axial and shear test (after Bawden et al., 1994)
Figure 2.2.6 illustrates one configuration for direct-shear testing. In this
arrangement the direction of motion is always parallel to the separation plane as
would be the case on the basal plane of a sliding gravity block, for example (see
Figure 2.2.3). The orientation of the cable, in this setup can be varied from 90
degrees (perpendicular to the surface) to 135 degrees (so that the cable is axially
pulled in tension as well as sheared) to 45 degrees (the cable must first kink in
compression before shearing). In this test, free dilation (aperture increase) of the
sliding plane is prevented.
Figure 2.2.7 shows a fundamentally different type of shear test. In this test, the
cable is always perpendicular to the separation plane. This is analogous to a
cablebolt installed perpendicular to a laminated hangingwall. The testing frame
allows for separation (aperture increase) to occur in addition to shear at numerous
angles with respect to the plane (and to the cable). For example an angle of 45
degrees would represent a slab falling straight down from an inclined, laminated
wall, inclined at 45 degrees. Bawden et al. (1994) describe some of the many
important procedural details required for successful testing of this kind. In
particular, it is essential to provide the appropriate confining boundary conditions
to the grout column since this is a key parameter controlling shear behaviour.

Figure 2.3.1: The cablebolt system (angles and spacings are examples only)
2.3 The Cablebolt System
Figure 2.3.1 illustrates the makeup of the cablebolt system which is comprised
of the cablebolt array and the cablebolt element itself. The overall performance of
the cablebolt system is the result of a complex relationship of these components
and of their even more complex interaction with the rockmass.
2.3.1 The Cablebolt Element
The individual cablebolt element is made up of several mandatory and several
optional components.
The Cable Tendon
The steel strand, the set of paired or multiple strands or, in recent
developments, the fibreglass wire cluster makes up the cable tendon. Most of this
book deals with standard and modified (flared) cable configurations based on the
7-wire steel strand.
Grout
The grout forms the link between the cable and the rock mass. Chemical grouts
for cablebolting have undergone some experimental use. However, this book will
be primarily concerned with cement grouts.
Borehole
Some cablebolt elements are sensitive to the condition and diameter of the
borehole and the properties of the rock surrounding the borehole.
Interface Mechanics
The mechanics of the interface between the cable and the grout usually
determine the overall behaviour of the system. These mechanics are described in
detail for the plain strand cable. The overall system capacity can be limited by the
efficiency of the bond strength of the cable-grout interface which can be extremely
sensitive to quality control, rockmass stiffness and rock stress change after
installation. The various developments in modified cable geometries are primarily
aimed at changing the mechanics of load transfer at this interface.
Surface Fixtures and Restraint Elements
Plates, barrel and wedge assemblies, and surface retention elements such as
mesh and straps are important aspects of the cablebolt element. Economically
practical cable spacings may not always be sufficiently tight to contain smaller
surface blocks and wedges. Surface fixtures must perform this role.
Tensioning
The degree (or absence) of tensioning (pre- or post-grouting) can have an
influence on the performance of the cablebolt system in fractured ground.

2.3.2 The Cablebolt Array
The behaviour of a cablebolt system is determined by the make up of the
individual cablebolt element and by the cablebolt array as a whole. The
components of the array are as follows:
Spacing or Density
A denser cablebolt pattern (a smaller spacing) will have a higher overall
stiffness and load capacity, but will be subject to economic constraints. The
ultimate ductility will not be substantially affected.
Face Pattern
Dominant joint orientations and block shapes may necessitate a particular
geometric pattern in order to intersect as many free blocks as possible.
Compatibility with face restraint systems such as straps may also dictate a
particular pattern. Normally, for a constant density, small changes in spatial
pattern will have a minimal influence on performance.
Length
The length of the cablebolts in an array should be determined by considering
the required and actual capacity of the system and the height or thickness of any
discrete geometric feature (i.e. wedge or broken zone) being supported. (For
example, 20 metre cablebolts in a gravity-based design are unnecessary if the
capacity of the system is equivalent to only 4 metres of rock).
Orientation
The orientation of an array is important when trying to optimize support
efficiency in cases where the directions of loading and induced displacement
(controlled by stress, excavation geometry, gravity and structural discontinuities)
are known. Ideally, cablebolts designed for holding (gravity loading) should be
aligned along the direction of displacement. Where applicable, to optimize
reinforcement efficiency along shear planes (Section 2.8.3), cables should be
oriented at 20-- 40- to a sliding surface (positioned to induce tension in the bolt).
Sequence and Timing
Timing of support installation and sequence with respect to mining can have an
influence on the system performance by altering the displacements sustained by
the system, the pre-cabling deterioration of the rockmass and the stress change and
capacity change experienced by the cable system (Section 2.6.2).
2.4 7-wire Steel Strand
Modern cablebolts for hard rock mining applications are primarily based on the
7-wire steel strand originally manufactured for use in prestressing concrete in civil
construction. Other configurations include wire rope, fibreglass tendon, and other
polymer and composites.
Most of this chapter will be concerned with variations on the 7-wire steel strand
cable. Before proceeding, a few definitions regarding the makeup of the steel
strand cablebolt are required:
2.4.1 Definitions
Cablebolt element
A single, complete in-hole assembly, including all of the contents of the
borehole (cable, grout, etc.), surface fixtures, the borehole specifications and
properties, as well as the face restraint system.
Cablebolt
The steel component contained within the borehole (beyond the collar). This
includes the steel strand or set of multiple strands in a single borehole and any
modifications (bulbing, birdcaging, nutcaging, ferruling, buttons, etc).
Strand
Any length of finished material which comprises a number of wires (i.e. six)
spun together in helical form around a centre wire as in seven-wire steel strand
(six wires around a centre wire).
Wire
A single continuous length of steel, round in cross-section.
Inner or king wire
The centre wire in a strand which is straight.
Outer or helical wires
Six wires are wound around the centre or king wire and are heat treated (stress
relieved) to form a continuous helical spiral.
Lay
The direction of the wire described as left or right hand lay. This has no effect
on performance in the field but is a concern in testing when designing anti-rotation
devices and when coiling cables for shipment.

Figure 2.4.1: Geometrical properties of seven-wire strand
Lay length
The axial distance along the strand required for an individual outer wire to
return to its original radial position; also called the pitch length.
Flutes
The V-shaped helical grooves created along the strand as the six outer wires are
wound around the king wire.
Nominal diameter
The diameter assigned in specifications for wire or strand. In the case of the
strand this will be measured from the outermost extents of the cable cross-section
though the centre of the king wire.
Nominal area of strand
The sum of the cross-sectional area of the individual wires.
Calculated mass of strand
The mass per length calculated from the nominal area and the unit weight of
steel (taken as 7850 kg/m ).3
Gauge length
The length over which deformation and strain is measured in a test. Usually the
central third or the complete portion of the total free length used in the test.
Coil Diameter
Steel cable can be shipped in a continuous length wrapped into a large coil. The
inside diameter of this coil must be greater than a specified minimum to avoid
disruption of the strand integrity (kinking and/or unravelling).
Figure 2.4.2: Manufacture of seven-wire, stress-relieved reinforcing strand
2.4.2 Strand Construction
Manufacture
The individual wires are manufactured from round wires, cold drawn from hot-
rolled high carbon steel wire rod which has been treated to make it suitable for
drawing. The steel should not have more than 0.060 % of sulphur and not more
than 0.060 % phosphorus (AS1311 - 1987).
The wires may be plain with mill finish or indented. This indentation is to
improve adhesive bond in concrete construction but is of dubious benefit in
cablebolting applications.
6 wires are laid helically around a slightly larger (1.02 - 1.03 times larger)
centre wire with a lay-length of 12 to 16 times the diameter of the final strand.
After stranding, the strand is stress relieved by continuous heat treatment. This
ensures that the helical wires maintain their form and reduces the time dependent
relaxation or creep of the strand under load (Collins and Mitchell, 1991). The
strand is then formed into coils (of greater than a specified minimum diameter to
ensure cablebolt integrity) for transport and sale.
The strand may be subsequently drawn and compacted, resulting in a reduced
cross-section and increased effective density of steel within the strand cross-
section. This process also results in a smoother surface profile as the outer strands
become flattened. While compacted or drawn cable has a higher tensile strength,
it has been used in mining with little measurable benefit and is not advised for
cablebolting due to the reduced mechanical interlock with the grout.

Figure 2.4.3: Variation in pullout performance under identical conditions for strands
from six different manufacturers (after Bawden et al., 1995)
Key Quality Indicators
Some important properties of the strand which can be tested by inspection at
the mine site in order of importance are:
+ The strand should not carry on its surface or between the wires any lubricant,
oil, rust or matter likely to impair its adhesion and friction with the grout.
+ The wires should show no tendency to unravel when the strand is cut.
+ The finished strand should be of uniform diameter (± 0.4 mm), of standard lay
length or pitch and should be free from kinks (plain strand). The centre wire
must be held tightly in place and show no signs of looseness.
+ Drawn cables with flattened outer wires should not be used.
+ The wires must be continuous with no more than one individual weld per 50
metre length of completed strand. The individual wires should be free of
defects such as splits and surface flaws.
If any of these requirements are not met, it indicates a deviation from the
accepted construction standards, and the strand should be returned to the
manufacturer. Such deviations from the standards have been shown to result in
serious degradations in pullout performance (Bawden et al., 1995) and steel
strength (Thompson, A.G., 1993, pers. comm).
Strand Construction Standards
The construction specifications for the 7-wire strand used in cablebolting are
covered by the following standards (or equivalent) relating to pre-stressing strand
for concrete construction:
North America (ASTM) A416-80
Australia (Australian Standards) AS1311 - 1987
Even though these standards refer to application in reinforced concrete, they
detail the necessary construction specifications to ensure optimum performance
of the strand in mining applications. While several grades of steel are covered by
these standards, the most commonly used product for cablebolt applications are:
Table 2.4.1: Standards for seven-wire strand (used for cablebolting)
ASTM A416-80 AS1311 - 1987
250 Regular Strand (not recommended)
* 270
* Super Grade (normal or low relaxation)
Extra-High Tensile Grade (normal or low relaxation)
* Most commonly used for cablebolting applications
Some basic geometrical specifications are as follows:
Table 2.4.2: Geometrical specifications for 270 - Super / EHT Grade strand
Specifications
(approximate - check
standards listed above)
Nominal Strand Diameter
(from crest to crest of opposing wires)
12.7mm
(0.5 )
15.2mm *
(0.6)
18mm
(0.7)
Nominal Area
mm
99 - 100 140 - 143 199
Lay-Length
mm
155 - 200 185 - 240 220 - 285
Calculated Mass
kg / metre
0.775 - 0.785 1.102 - 1.125 1.560
Minimum Internal Coil Diameter
mm
750 750 750
* NOTE: All of the discussion in this handbook involves 15.2mm strand cablebolts.
Do not use the design recommendations in this manual for other strand
sizes without applying an appropriate adjustment.

2.4.3 Strand Performance
Definitions
The axial performance of 7-wire steel strand is described by the following
specifications (There are no standards for shear performance):
Breaking Load
The tensile load applied to a length of strand at the instant of rupture.
Proportional Limit
The load or strain at which the elastic behaviour of the cable deviates from
linear.
Yield Strength and Proof Load
The yield strength is the tensile load applied to a length of strand to achieve a
non-proportional (inelastic) strain of 0.2 %. The proof load (also called yield
strength in ASTM standards) is similar to yield strength but is specified at a total
extensional strain (elastic + inelastic) of 1.0 %.
Elongation:
The total strain ( L / L) of a minimum specified gauge length of strand at the
instant of rupture of one or more wires in the strand.
Elastic Modulus
The theoretical stress (load/area) applied to a unit cross-sectional area to
achieve a unit elastic strain (100 %). The elastic modulus of carbon steel is 205-
210 GPa. The modulus of the 7-wire steel strand (based on the nominal area of the
strand) is somewhat less than this value due to the behaviour of the helical outer
wires.
Elastic Stiffness and Normalized Elastic Stiffness
The elastic stiffness is simply the slope of a load / deformation curve and has
units of MN/m. The normalized elastic stiffness is the stiffness of a unit length of
material or load / strain. It is quoted in units of MN / (m/m) or MN.
Relaxation
Relaxation is defined as the reduction in load with time of a specified minimum
length of strand held at a constant strain or elongation. This strain is achieved at
an initial load which is specified as a standard percentage of breaking load.
Relaxation is usually specified as percentage relaxation (drop in load as percentage
of initial load) after a standard time interval. Relaxation is directly related to creep.
Creep is defined as a rate of extension of a sample held at constant load.
Figure 2.4.4: Minimum strand performance specifications for cablebolting applications
Strand Performance Standards
Table 2.4.3: Approximate performance specifications for 7-wire steel strand
Specification
(AS 1311 - 1987,
ASTM A416-80)
See
Section 2.4.3
Figure 2.4.4.
Gauge *
Length:
mm or
x LL
(Lay-
length)
Regular /
250
Grade
Super Grade Steel Strand E H T /
270
Grade
Nominal Strand Diameter, mm
15.2 12.7 15.2 18.0 15.2
Breaking Load:
Min. (kN)
600 or
3 x LL
230 / 240 184 250 338 261 / 261
Yield or Proof Load
(kN) @ 1% Total
Strain (elongation)
200 or
1 x LL
196 / 204 156 212 287 222 / 222
Elongation :
Min. (Strain)
600 or
3 x LL
3.5 % (strain)
@ Breaking Load
Approx. Elastic
Modulus (GPa)
200 or
1 x LL
195 - 200 **
Below Proportional Limit
Normalized Elastic
Stiffness (MN)
200 or
1 x LL
26 - 27 19 - 20 27 - 28 39 -40 27 - 28
Relaxation
Max. @ 1000hr
80 % Brk Load
600 or
3 x LL
LOW Relaxation Strand: 3.5 %
NORMAL Relaxation Strand: 12 %
* All tests should have a minimum of 600 mm free length between end grips.
** Stiffness and modulus shown are for rotationally constrained case and with respect to nominal
area. When free rotation (untwisting) is permitted, the initial stiffness is reduced by up to 25 %,
rising to the constrained stiffness (tangent) with increasing strain. (Costello et al., 1976)

2.4.4 Strand Capacity Considerations
Cablebolt design tools such as rockmass classification, specialized empirical
tools such as the Mathews/Potvin Method, and mechanistic analysis provide
recommendations for spacing (or density) and length of cablebolts in regular
patterns. Spacing recommendations are normally based solely on the tensile load
bearing capacity of the strand while length specifications assure the ability of the
cablebolt to traverse the reinforced zone and penetrate into stable rock.
NOTE: Most bolting system design tools assume perfect load transfer (very
high bond strength), between the cablebolt and the surrounding rockmass,
which can be a very erroneous assumption as discussed in Section 2.6. This
capacity can be seriously impaired by ineffective bonding (interface shear
strength) between the cable and the grout. These issues are discussed later in
Section 2.6. In addition the ability of the cablebolt to withstand direct shear
(guillotining) is less predictable than the tensile strength. It can be assumed that
a 5 - 20 % reduction in capacity is possible in the field due to partial shear
loading, although this is heavily dependent on the loading angle (tension + shear).
The stiffnesses listed in Table 2.4.3 and illustrated in Figure 2.4.4 refer to the
performance of the steel strand only. The bond mechanics again govern the actual
observed stiffness and displacement capacity of a grouted cablebolt in the field.
This stiffness is relevant only when known debonded sections are present. In
addition, the shear stiffness of cablebolts is significantly less than the response to
tensile loading (Windsor, 1992; Bawden et al., 1994) and is extremely dependent
on loading angle (Section 2.8).
Under ideal conditions, over a finite area of supported surface, load capacity
and stiffness are directly related to the number of cablebolts installed. Doubling
the strand density (number of strands per unit area), either by reducing the cable
spacing (e.g. from 2mx2m to 1.4mx1.4m) or by using double strand cablebolts
(two strands in one hole), normally results in a doubling of both the load capacity
and the stiffness of the cablebolt system, although care must be taken when
grouting double strand cables to ensure full encapsulation of both strands.
Double strand cablebolts with spacings equal or greater than those for single
strand patterns do not improve the ability of the system to retain surface blocks or
fragments between the cables. This failure mode must always be considered
separately, particularly for cables on wide spacings (1.5m). Surface restraint
(rockbolts, screen) may be a necessary supplement to the cablebolt system.
Plating exposed and accessible cablebolts is always advisable if timing and
economics permit. In addition to reducing the sensitivities (of plain strand in
particular) to the influences of rock stiffness, stress change and quality control,
plates also provide another measure of surface retention and safety.
2.4.5 Corrosion of Steel Strand
Corrosion of high carbon steel strand can be a serious problem in long term
civil engineering applications. In mining, however, the incidences of cablebolt
corrosion causing serious problems are rare. This is due primarily to the short time
frame involved in open stope support in underground mining.
Corrosion problems observed by the authors in mining environments were
typically in long term support in open pits where the groundwater was acidic or
saline and in long term support in underground sulphide deposits. Cut and fill
applications in wet conditions where fractured stope backs could remain
(supported) for up to a year were notably susceptible to corrosion. Serious failure,
due to corrosion and rupture of the strand, can occur in such applications.
The nature of corrosion is extremely complex and a fundamental discussion is
beyond the scope of this book. It is the intent here to discuss some of the important
factors involved in corrosion so that the engineer may assess the potential for
problematic corrosion and take steps to prevent it or make the appropriate design
allowances for it.
Most common refined metals are inherently unstable ionic materials composed
of arrays of single atoms which possess a full compliment of electrons. Metals
such as iron normally tend to give up electrons at room temperature (gold is a
notable exception) and become involved in reactions leading to the formation of
more stable compounds such as iron oxide or iron hydroxide (rust). The release
of electrons is termed an anodic reaction and the acceptance of electrons a
cathodic reaction. Both reactions must occur for corrosion to take place. Since
metals such as the iron found in steel cable are normally willing to give up their
electrons, it is normally the presence of a cathode which determines the corrosion
potential.
The cathodic reaction (involving the consumption of electrons released
anodically from the iron) can be made possible by the presence of an acid,
sulphate, water and/or oxygen.
Corrosion of steel (iron) can be divided into four basic categories (Illston et al.,
1979; Pohlman, 1987):
+ Dry corrosion
+ Wet corrosion
+ Corrosion of immersed metals and alloys
+ Induced or accelerated corrosion (includes influence of stress)
The following discussion is confined to corrosion of cablebolts and as such is
incomplete as a comprehensive examination of general corrosion.

Dry Corrosion
Dry corrosion is an inevitable consequence of medium- to long-term storage of
cablebolts in even the most ideal conditions. It involves the formation of iron
oxide (Fe0) as iron atoms combine with atmospheric oxygen. Once the process
initiates on a clean surface, it spreads fairly rapidly to involve most of the exposed
surface. While Fe0 forms an adherent film on steel surfaces and can actually form
an impervious layer, it can be vulnerable to cracking and as such fresh iron is
constantly being exposed and the process continues. In the perspective of
cablebolting in mining, however, dry oxidation is a relatively slow chemical
process and is of only minor consequence. Light surface (dry) corrosion has been
shown (Goris, 1990) to improve bond performance of cablebolts by up to 20% in
ideal conditions, although deliberate rusting of cablebolts is not advocated by
the authors. The process is accelerated by higher surface temperatures (e.g. if the
cables are exposed daily, over long periods, to direct and intense sunlight).
Heavy surface rust on newly shipped cables is usually the result of exposure to
moisture and subsequent atmospheric corrosion which can be very detrimental to
the performance of the cablebolts.
Wet or Atmospheric Corrosion
In a wet or humid environment, the corrosion process is accelerated and can
involve a wider variety of cathodic reactions. Water and oxygen become jointly
involved in the cathodic reaction and result in other compounds such as 2Fe(OH) ,3
Fe O (magnetite), or Fe O (hematite). These compounds are much less adhesive3 4 2 3
then FeO and less likely to form a self-arresting film.
Corrosion products formed on cablebolts by wet corrosion are more likely to
have a greasy feel as compared to the dry, rough texture of FeO film and are more
likely to be associated with other film substances such as oils and additional
moisture. These products are likely to have a detrimental effect on bond capacity
of cablebolts. Clearly, unchecked corrosion reduces the cross-sectional area of
steel in the cable and ultimately reduces the tensile capacity of the steel to
unacceptable levels. Ductility and displacement capacity is also reduced
(embrittlement).
The presence of water on the surface of the cablebolt also increases the
potential for galvanic corrosion. The same wet corrosion cathodic reactions occur,
accelerated by the presence of an electrolyte such as chloride, sulphate or
hydroxide. Without electrolytes in a static solution, the corrosion process is self-
limiting. Iron ions (e.g. Fe ) move into solution adjacent to the steel surface2+
leaving behind free electrons (2e) in the steel solid. The concentration of iron ions-
in solution and free electrons in the steel creates an electrical potential difference
which resists further dissolution of iron ions.
Figure 2.4.5: The mechanisms of rusting of mild steel (after Illston et al., 1979)
The effects of electrolytes in the surface water is best illustrated in the above
example. A drop of water on the surface of the steel contains a dissolved
electrolyte such as sodium chloride (which forms a solution of free sodium, Na ,+
and chloride, Cl, ions). The presence of electrolytes permits the transport of iron-
ions as FeCl away from the corrosion (anode) site at the centre of the drop. At the
same time, water and oxygen combine at the perimeter of the drop with the free
electrons from the steel to form hydroxide ions (OH ) balanced by Na in solution.- +
These move in the opposite direction to the FeCl generating a current (electron
flow) in the steel supplying electrons to the drop perimeter as more iron ions go
into the solution at the drop centre. Between the active centre (anode) and the drop
perimeter (cathode) the iron ions combine with the hydroxide to form ferrous
hydroxide. This in turn becomes a relatively stable and complex hydrated oxide
known as rust. The sodium and chloride transport ions are freed to carry on the
process. The cyclic nature of the process combined with the fact that the corrosion
product (rust) is not deposited at the anode (as it is with dry corrosion) means that
this form of galvanic corrosion is not self-limiting and can be very aggressive.
This is particularly true in mining environments given the high concentration of
chloride and sulphate ions in mine waters (Minick and Olson, 1987).
Moist corrosion is particularly enhanced by crevices such as those formed by
the flutes of a cable. Crevices are particularly good at retaining moisture and the
conditions are perfect for differential aeration with low oxygen supply at the tip
of the crevice compared with the rest of the cable. If a weak electrolyte is present,
an aggressive corrosion cell is thus generated. This corrosion is particularly
detrimental as the corrosion product (rust) readily fills the flutes of the cable
preventing the penetration of grout and seriously reducing the cable/grout
interlock essential for cable bond strength.

Figure 2.4.6: Corrosion rates for immersed carbon steel (after Bryson, 1987)
Immersion of Metals and Alloys
It is the differential electrical potential between the anode (+) and the cathode
(-) which is key to the moist corrosion example described above. This differential
is primarily generated by the difference in oxygen availability between the edge
and the centre of the water droplet.
Differential potentials can also be generated by the presence (and contact) of
dissimilar metals immersed in an oxygenated electrolyte solution (Illston et al.,
1979; Bryson, 1987). Corrosion induced by such a coupling can be extremely
aggressive and can result from the designed use of dissimilar metals (steel cables
with aluminum plates or anchors) or from the presence of cablebolts in a rich
sulphide ore. Indeed, rock bolts in sulphide ore bodies have significantly reduced
service lives (Hoey and Dingley, 1971; Gunasekera, 1992).
Corrosion cells can also be generated on cablebolt surfaces at the point where
abrupt transitions in environment occur. These include differential grout coverage,
for example, at the borehole collar, at penetrating cracks in the grout, where the
cable crosses a local water table, or within voids in the grout column.
Oxygen (atmospheric or dissolved) is the critical component of the cathodic
reaction discussed so far. The concentration of oxygen is therefore a critical factor
governing the rate of corrosion. In aqueous environments with high levels of
acidity or low pH, however, the hydrogen (H ) ions in the acid solution react+
cathodically with the free electrons in the steel to form hydrogen gas (H ). This2
reaction is countered as before by the release of iron ions from the steel and does
not require the presence of oxygen. While oxygen concentration normally controls
corrosion rate (loss of iron ions), the acid (H ) reaction dominates below a pH of+
4 and can become extremely aggressive (Figure 2.4.6).
Although it is not as common as oxygen related corrosion, acid corrosion can
pose a serious hazard to mine support (Gunasekera, 1992) due to its accelerated
rate. Sampling of groundwater and/or mine water for pH is relatively simple so the
risk can be easily determined. In Canada, mine water with a pH of 2.8 has been
recorded in underground mines, and measurements of 3-4 are not uncommon
(Minick and Olson, 1987). Acidic mine water can often be linked to the oxidation
of sulphide ores (primarily pyrite and marcasite) resulting in the generation of
sulphuric acid and pH levels as low as 1.5-2 (Gunasekera, 1992).
In addition, there are many species of bacteria which flourish in the
underground environment and which greatly accelerate the breakdown of
sulphides to form sulphuric acid. Different species are active with and without the
presence of oxygen. Such bacteria can accelerate the production of acid in mine
waters by a factor of four with a related increase in corrosion rate.
Accelerated Corrosion
Of primary consideration in cablebolting is the acceleration of any of these
corrosion processes at points of excessive strain in the cablebolt. As steel is
strained in tension or in shear across a joint in the rock by rockmass movement,
or bent by improper plate installation, the susceptibility to all forms of corrosion
increases. Any protective surface rust is cracked by such strain exposing fresh
surfaces. Microscopic cracks formed in areas of high strain create corrosion
conduits beyond the steel surface. In addition, the strained ionic bonding in the
metal increases the potential for iron-electrolyte interaction and hydrogen
embrittlement (Littlejohn and Bruce, 1975).
This so-called stress corrosion cracking is important because cables will tend
to corrode much more rapidly in aggressive environments exactly when and where
their mechanical integrity is most tested and is most critical. In the case of grouted
cablebolts, load concentrations along the cable length are usually related to full
cracking and separation across the grout column. This allows direct and focussed
attack on the stressed steel by corrosive agents. Stress corrosion is often the final
mechanism in cablebolt failure in corrosive environments.
Cablebolt Geometry Effects
In general, the high carbon steels used in the manufacture of cablebolt strand
are more corrosion resistant than the steels used in conventional rock bolts.
Nevertheless, certain features of the grouted cablebolt which increase its potential
for detrimental corrosion include the presence of flutes (v-grooves), internal
channels between the outer wires and the king wires, as well as the formation of
concentrated corrosion sites at separation planes in the rock and grout. Voids and
bubbles in the grout column also create potential corrosion cells.

Summary Recommendations for Corrosive Environments
Corrosion is rarely a problem in open stope cable support, simply due to the
short service life involved. Cut and fill stopes can be open for up to a year or more
and overhead cables should, therefore, not be allowed to corrode to unacceptable
levels during this time. Fractured, sulphide ore bodies require special attention in
this regard. Corrosion of cablebolts (and other steel support) in permanent mine
openings can cause serious problems in terms of safety and rehabilitation. In
addition to normal capacity reduction, corroded cables tend to become brittle and
can suffer reduced effectiveness in dynamic loading situations. The factors which
contribute to corrosion are often complex, are compounded in an underground
environment, and are very difficult to combat in areas of high severity.
Nevertheless, the following is a brief list of remedial measures for use when
corrosion has been identified as a problem (Littlejohn, 1990; Gunasekera, 1992).
Cablebolt storage
+ Store cablebolts in a dry location, preferably moving them underground to the
working site only when required. Long-term storage outside, under the sun or
exposed to the elements should also be avoided.
+ Do not allow water to collect on the cablebolts. Corrosion will quickly fill the
flutes reducing bond strength and potentially pitting the steel.
Installed cablebolts
+ High humidity accelerates corrosion. Good ventilation at all times can help to
reduce this factor.
+ Use caution when installing cables in areas with flowing water.
+ Avoid any use of cements, mixing water or admixtures containing chlorides,
sulphides or sulphites.
+ Grout voids and bubbles increase corrosion potential.
+ Request that plates, barrels and wedges, and other fixtures are electro-
chemically compatible with the high strength carbon steel used in strand.
+ Long rust stalactites growing rapidly from the ends of uphole cables indicates
potentially severe strand corrosion up the hole.
+ Sulphate resistant grouts are alkaline and can counteract acidic mine waters.
The use of this cement does not permit the use of such waters for grout mixing.
Severe corrosion
+ Epoxy-encapsulated cables are available for use in corrosive environments
(Windsor, 1992). Note that such coatings may not be resistant to all forms of
corrosion and that the coating must penetrate the strand, encapsulating the king-
wire to prevent focussed corrosion down the centre of the strand.
+ Galvanized cable would be of use against non-acidic corrosion.
+ Grease can protect ungrouted lengths of cable (at the collar, for example).
+ Other more costly measures such as cathodic protection are discussed in
Littlejohn and Bruce (1975) and Littlejohn (1990; 1993).
2.5 Grout
Cablebolts are classed as continuously coupled, friction anchored devices. This
means that there is a continuous coupling between the cable and the rock mass.
This is achieved through the use of grout. While this handbook focusses on
Portland cement grouts, there are several other material options:
Resins
Other more rigid grouted systems such as grouted rebar, can be effectively
grouted with epoxy-resin grouts. Other chemical products such as urethane and
other polymers have also seen limited use. The main problem with these chemical
grouts when used with cablebolts seems to be with installation. It is difficult to
spin a cable into a hole full of grout cartridges as is done with rigid rebar.
Pumping and placing chemical grouts is often an unpleasant experience in an
underground environment and may conflict with local safety regulations. Progress
is being made, however, in these areas (Goris and Tadolini, 1993).
Shotcrete
The use of shotcrete for cablebolt grouting would have advantages in mines
already using shotcrete and would reduce equipment requirements where the two
support systems are used together. Hassani and Rajaie (1990) have pull-tested
shotcrete grouted cablebolts in the laboratory and have obtained favourable results
showing slightly reduced initial bond stiffness followed by comparable load
capacity and enhanced residual strength at large displacements.
Portland Cement
Portland cement grout without aggregate remains the primary grout material.
The design of the grout for cablebolt applications is often critical to success of the
operation. Proper grout design is based on the following considerations:
Installation considerations:
+ Installation method
+ Cablebolt geometry (type)
+ Mixing time
+ Pumping rate, efficiency and limits
+ Initial setting time
Performance considerations:
+ Cablebolt geometry (type)
+ Bond strength and load transfer mechanics
+ Curing time
+ Grout performance with time
+ Environmental sensitivity (corrosion, etc.)
These requirements can be mutually contradictory in certain circumstances. It
is important to reach an adequate compromise when designing a grout mix.

Figure 2.5.1: Compressive strength development in pastes of pure cement compounds
(after Hyett et al., 1992 and Mindess Young, 1981)
2.5.1 Composition of Cement Grout
The most common form of grout used in cablebolting, and indeed in over 90%
of grouting and concrete operations (C.P.C.A., 1984), is ordinary Portland
Cement. It is made from limestone and clay or other suitable materials, heated
together and pulverized to form a powder (particle size in the range of 0 to
0.05mm) which is rich in calcium silicates (C.C.A., 1968; Mindess and Young,
1981; Hyett et al., 1992). Small amounts of gypsum (calcium sulphate) are also
added to moderate the behaviour of the final product when mixed with water.
The final product contains approximately 45-60% C S, 15-30% C S, 6-12%3 2
C A, and finally 3-8% C AF, where C = CaO (lime), S = SiO (silica), A = Al O3 4 2 2 3
and F = Fe O (Ferric Oxide). The relative content of these minerals controls the2 3
behaviour of the cement product (Mehta, 1986; Hyett et al., 1992) when mixed
with water and allowed to hydrate. This hydration is a chemical reaction which
evolves heat and forms new compounds and ultimately a strong solid mass. It is
not a drying process and it is not reversible (it is impossible to reuse hydrated
clumps of cement). Figure 2.5.1 below illustrates the contributions to strength and
curing rate of the individual cement minerals. This gives an indication of the
effects of varying the proportions of these minerals.
The grain size is equally important in determining behaviour. The rate of
cement hydration (reaction with water) is a function of the blaine or total particle
surface area within a unit mass of cement. The greater the surface area per volume
of cement (i.e. the finer the grout), the faster the reaction (set).
2.5.2 Varieties of Portland Cement
There are several types of cement which are applicable for use in cablebolting.
They are all variations of normal or ordinary Portland cement. They contain the
same basic minerals but in different proportions. There may be other acceptable
variations which utilize various combinations of fly ash, blast furnace slag and
other materials, although testing for strength and performance in the operating
environment should be carried out. Most of the grout related discussions in this
book focus on ordinary or normal Portland cement.
Ordinary or Normal Portland Cement
This is the most common and least expensive material for grouting. Only in
special circumstances should it be necessary to use a different grade of cement.
Typical concentrations of cement components in normal Portland cements such
as the Canadian specification Type 10 are approximately 50% C S, 25% C S, 12%3 2
C A and 8% C AF. The typical blaine of normal Portland cement is approximately3 4
350 m /kg. This value can vary between manufacturers and accounts in large part2
for the variation in performance of otherwise identical batches of cement
(Cortolezzis, 1991; Hyett, Bawden and Coulson, 1992).
Rapid-Hardening or High-Early Portland Cement
High-early cement (Canadian Type 30) contains a greater proportion of C S3
relative to C S and a smaller grain size (and therefore higher blaine - 450 m /kg).2
2
This gives it a higher early strength. The final strength (after 28 days) will not be
greater than ordinary Portland cement, however, and in fact may be slightly less
due to shrinkage and embrittlement. The finer grain size and higher initial
reactivity may also lead to premature set and lumpy grout during storage in a
humid mine environment. It is advisable to perform bond testing (Section 2.2.2)
at lower water:cement ratios (0.35) to ensure that grout disintegration due to
early heat generation and expansion is not a problem. Such difficulties have been
reported (Oliver, 1995, personal communication).
Portland Blast Furnace Cement
This cement is made by combining normal Portland cement with suitable
granulated blast furnace slag. Concrete standards permit the maximum proportion
of blast furnace slag (by weight) of 65%. This cement can be used in cablebolting
with no appreciable reduction in cement quality after hydration. It may tend to
gain strength more slowly and evolve less heat during hydration. In concrete
applications it is often claimed to be more resistant to chemical attack, although
this has not been investigated in cablebolting applications. The primary advantage
to this type of cement is its reduced cost in regions where blast furnace slag is
readily available.

Sulphate-Resistant Portland Cement
This cement contains a lower proportion of C A, and a higher content of C AF.3 4
The reduction in the former increases this cement's resistance to attack by sulphate
compounds which can be abundant in some mine waters. Sulphate attack is a form
of corrosion and can seriously degrade concrete strength with time. In short term
applications, sulphate resistant cement would be required when the sulphate
concentration in the groundwater exceeds 2500 parts per million. Greater than
1200 parts per million of SO would suggest its use in long term installations3
(C.C.A., 1968). It is no more resistant, however, than normal Portland cement to
acids or other dissolved salts which attack cement. Sulphate resistant cement is
also suitable for use in high temperature environments at depth in mining where
normal Portland cement may be predisposed to excessive cracking and shrinkage.
Special low-heat cements are also available (C.P.C.A., 1984) but these are usually
expensive and of little advantage for cablebolting.
Low-Heat (Portland and Blast Furnace Slag) Cements
There has been some use of low-heat Portland cement and low-heat blast
furnace slag Portland cement in cablebolting applications where the combination
of internal rock temperature and the heat evolved during the hydration of cement
has been thought to cause temperature related problems such as cement cracking
and shrinkage. It is unclear whether this is a real problem in the temperature
ranges encountered in rock. White, powdery grout in the borehole is an indication
of excess heat generation. Sulphate resistant cements and regular blast furnace slag
cements both evolve heat more slowly than normal Portland cements and are an
adequate and more economical option in high temperature environments.
Silica Fume Portland Cement
The addition of silica fume, an extremely fine (smoke-sized) particulate by-
product of the production of silicon metals, results in a higher strength and faster
setting grout mix. The extremely fine particles and the high glass content of silica
fume creates an enhanced hydration reaction, increasing strength and reducing
porosity. As a result, this cement can provide higher cablebolt bond strengths
(Hassani et al., 1992) and improved protection in chemically aggressive
environments.
Sanded Cements
Some research has involved the addition of sand to the cement mix (Goris,
1990; Gendron et al., 1992). This increases the base strength and stiffness of the
grout. It is unclear, however, whether this increases the performance (i.e. bond
strength) of grouted cables.
2.5.3 Care and Quality of Cement and Water
Anhydrous Portland Cement, (Dry Powder)
The prime concern with respect to storage and transport of grout is pre-
hydration. Due to moisture influx, the cement powder will begin to set (hydrate)
and take on a hard or lumpy texture. Portland cement cannot be reconstituted once
it begins to hydrate. Lumps will remain in the grout and adversely affect the
strength and stiffness within the grouted borehole column and will therefore
reduce cablebolt performance. In addition, lumpy grout can cause numerous
problems during installation of the cablebolt. The following guidelines should be
observed:
+ 'Rapid-hardening' or `high-early' cement is particularly susceptible to pre-
hydration and should not be used in mining applications unless maximum cable
capacity is needed in less than 7 days from grouting.
+ Plastic lined bags are not recommended due to condensation (and hydration)
problems within the bags. Shrink-wrapped stacks of bags are preferable.
+ Reduce the storage time on site - Order small!...order often!
+ Upon arrival at the mine site, the bags should be checked. If hard bags are
detected, the whole batch should be returned.
+ Store the bags above ground in a dry warehouse until ready for use.
+ Less than 3 days before use, transport underground on a raised flat and covered
with a waterproof tarpaulin. Multiple tarps should have generous overlap.
+ Do not stack bags higher than 1.5 m to avoid compaction and set.
+ When grout bags are being emptied into the mixer, any hydrated lumps should
be screened and discarded. Do not attempt to break up the lumps to fit through
the screen for use in the mix.
Water
Concrete standards require that water for cement production be potable
(drinkable). This will ensure purity standards required for adequate concrete
quality. In cablebolt grouting operations, this is also the optimum condition. If this
is not possible or practical (as is usually the case underground), then the following
guidelines should be observed:
+ Grout strength tests should be performed as described in the next section.
Cement mixed with mine water should achieve in excess of 90 % of the 28 day
strength of the cement mixed with distilled water. Set times should be
monitored and compared for acceptability.
+ The chloride content of the water should be low. Saline water will result in
limited deterioration of cement strength but will cause unacceptable corrosion
of cable steel.

W C
M kg
M kg
V litres
M kg
W
C
W
C
:
( )
( )
( )
( )
= =
W
V litres
V litres
V litres
V V litres
C
W
CP
W
W C
= =
+
( )
( )
( )
( )
W
W C
W C
W
C
W
C
W
C
C
C
=
+
×
=
+
1
1
ρ
ρ
ρ
ρ
ρ
:
; :
2.5.4 Properties of Fresh Cement Paste
When anhydrous cement is mixed with a quantity of water, the resultant
mixture is called fresh cement paste. This is the state in which the cement is
pumped into the borehole for cablebolt grouting. When the phrase pumping
grout is used in this and other literature, the reference is, of course, to the
pumping of fresh cement paste. This section defines some of the important
engineering properties of fresh cement paste.
Water:Cement Ratio (W:C)
This is the most important property of cement grout or paste. It is defined as the
ratio of the mass of water (M ) used to create the fresh cement paste mix, to theW
mass of anhydrous cement powder (M ) used in the same mix. Remembering thatC
the weight of water in kilograms is equivalent to the volume of water (V ) inW
litres;
Water Content (W )c
This is an alternative expression to describe the composition of fresh paste. This
is defined as the volumetric percentage of water with respect to V , the totalCP
volume of cement paste (V of anhydrous cement + V of water);C W
and
where , = 1.0 kg/l and , . 3.15 kg/l (Hyett et al., 1992) are the densities ofW C
water and anhydrous cement respectively.
Most of the discussion in this book refers to W:C and not to W . TheC
conversion is presented here for comparison with other literature.
Figure 2.5.2: Wet bulk density of fresh cement paste (after Hyett et al., 1992)
Wet Bulk Density (,, )c
Wet bulk density is the mass per unit volume of fresh cement paste (freshly
mixed and unset). This value is calculated by simply filling a container of known
volume (say 10 litres) with fresh cement and determining the mass of the full
container. Subtract the mass of the empty container from this volume. Divide the
remaining mass (mass of the cement alone) by the volume of the container (1 litre
= 0.001m ) to obtain the wet bulk density. There is a theoretical relationship3
between water:cement ratio, W:C, and the wet bulk density (Hyett et al., 1992):
W:C = 0.757 x ln (, - 1.0) - 0.333cp
This relationship is adequate below a wet density, , , of 2000kg/m (W:CCP
3
greater than 0.39). However, due to air entrainment in thicker grouts, the
relationship, which applies to fully saturated cement pastes, breaks down. Below
a W:C of 0.33, there may be insufficient water available to completely saturate the
cement grains (and achieve complete hydration) and/or air pockets can become
permanently trapped in the thick paste as it flows. This relationship is also valid
for hardened (hydrated) cement samples subject to the same constraints. The graph
below shows acceptable bounds to be used when calculating W:C from , , orCP
vice versa. Note that the data shows a wide scatter. This should be kept in mind
when using , as a quality control measure for W:C.CP

Mixing and Cement Saturation
Dry anhydrous cement particles are inert until moisture adsorbs (bonds to the
surface molecules) to the surface and initiates the formation of filament structures
which in turn interact with neighbouring particles to form the physical structure
of cement. This means that in order to achieve a uniform cement paste with
optimum properties, thorough mixing is required. A shear type mixer is one which
induces velocity differentials within the grout mix. This means that the grout is not
simply spun in the mixing bin but undergoes radial and axial convection (with
respect to the bin geometry). The higher the rate of shear, the more opportunity
the cement particles have to adsorb hydration water. The most common type of
shear mixer uses rotating blades to agitate the grout - water mixture. Key design
considerations for grout mixers and pumps are described in Section 2.11.
Viscosity of Fresh Cement Paste
Viscosity is a quantity used to express the resistance of a fluid to internal
dynamic shear - as occurs in mixing and pipe flow. Essentially, the higher the
viscosity, the greater the resistance to mixing, handling or flow.
A material which has a low viscosity such as water requires little effort to
transport at low to moderate pumping velocities (normally, internal shear increases
at higher velocities). The movement, through a constriction of pipe, of a highly
viscous material such as a thick grout, requires substantial pressure.
Plasticity of Fresh Cement Paste
Plasticity implies a resistance to static shear. Plastic grout pastes can maintain
form at rest (or remain, for example, in an unplugged uphole). Grouts of W:C less
than 0.375 exhibit a combination plastic-viscous or Bingham behaviour.
Hydraulic Behaviour of Fresh Cement Paste
Table 2.5.1 gives a brief definition of the various types of flow classifications
(I.S.R.M., 1991) for cement grouts used for cablebolting. This range of behaviour
explains, for example, the designation of W:C ranges for:
+ Toe-to-collar uphole grouting (W:C=0.3 to 0.35) where the grout must remain
in an unplugged uphole once placed.
+ The grouting of bulbed strand (W:C =0.4 to 0.45) where it is necessary to
compromise between the strength of a thicker grout and one which will
penetrate the bulbs.
+ Grout loss problems and associated hazards associated with pumping thin
grouts (W:C 0.4) in fractured ground. Thicker grouts (Bingham or
thixotropic) would slow and stall in a fracture.
Hydraulic Behaviour of Fresh Cement Paste (cont.)
Table 2.5.1: Key hydraulic classifications of grout flow
W:C Hydraulic Classification Conceptual Shear
Response
Pipe Flow
Velocity Profile
0.50 L
L
L
I
N
C
R
E
A
S
I
N
G
V
I
S
C
O
S
I
T
Y
L
L
L
Newtonian (Fluid) Flow
A linear relationship exists in
laminar (parallel) flow between
shear rate or velocity gradient
and shear stress as defined by
the viscosity constant, µ.
That is, faster pumping takes
proportionally more effort.
Pure water is an example of a
Newtonian fluid.
0.50
to
0.375
Pseudoplastic
As W:C decreases, the
relationship becomes non-linear.
Specifically, the spontaneous
viscosity is high when the fluid is
at rest and decreases with
increasing shear rate or velocity
gradient. As the mixing or
pumping rate increases, further
unit increases require smaller
increases in effort.
0.375
to
0.30
Bingham (plastic)
At low W:C ratios, the grout
behaves as a plastic solid at low
shear. As internal shear stress
increases beyond a threshold
value, J , the material behaveso
like a pseudoplastic fluid. The
grout returns to a semi-solid
(plug) as shear drops below J .o
Once the grout stops flowing or
mixing, it is difficult to restart.
0.45
to
0.30
with
admix-
tures
Thixotropic
This material can be either
pseudoplastic or Bingham in the
case of grout. (Thixotropic
behaviour is often confused with
Bingham). The term describes
behaviour in which viscosity
decreases with duration of
agitation (shearing). The
viscosity is restored to the
original value after a rest period.

Workability of Fresh Cement Paste
Closely related to the consistency, viscosity and hydraulic properties of grout
is its workability or behaviour during the placement process. There are four main
concerns with regards to workability which are often contradictory:
+ The grout must be fluid enough (low viscosity) to allow it to be pumped some
distance (typically 5m to 30m) along a grout tube and, where applicable along
a length of small diameter breather tube. It has already been seen that this
flowability and pumpability increases with increasing W:C ratio. W:C = 0.35
represents a lower limit for some commercial piston pumps while W:C = 0.3
has been successfully pumped and placed with more powerful progressing
cavity pumps and larger grout hose. It is doubtful that grouts of W:C0.35 can
be reliably pumped back down a small diameter breather tube ( 13mm I.D.).
+ The grout must be viscous or plastic enough to hold itself in a borehole against
the pull of gravity in a top-to-bottom installation. It must also resist flow into
open fissures intersecting the borehole. Viscosity and plasticity increase with
decreasing W:C ratio. Grouts of W:C 0.38 fulfil the first requirement and are
recommended for uphole installations where a breather tube and collar plug are
not used. Grouts up to W:C=0.40 can be used in fractured ground providing
that the cracks are not extreme in aperture.
+ The grout must fully encapsulate the cable and must therefore be fluid enough
to penetrate into the grooves of the steel strand cable or must fully penetrate the
cage or bulb of modified (flared) strand cables. The ability to fully cover the
cable surface and penetrate into the cable geometry increases with increasing
W:C ratio. Grouts of less than 0.35 are not recommended for flared cables such
as bulbed or birdcaged strand. Grouts with W:C below 0.35 may not be able to
reliably encapsulate the surface geometry of a regular strand cable.
+ The grout must resist initial set long enough to allow pumping and placement.
Initial set results from a physical attraction between wet cement particles. It
leads to an early stiffening of the paste and typically occurs between 10 and 60
minutes after initial mixing for the range of grouts used in cablebolting. This
should not be confused with hydration which leads to the ultimate strength of
the cement grout. A grout of W:C=0.3 will allow less than 15 minutes for
placement, while a grout with W:C=0.45 will give up to 45 minutes of
workability.
It will be shown in later sections that lower water:cement ratio grouts (0.35-0.4)
are recommended to optimize the frictional cable/grout bond strength of the
cablebolt. These thicker grouts can create slight problems in pumping and flow
which can be easily corrected without increasing the W:C ratio to unacceptably
high values. The following sections discuss these problems and their solution.
Figure 2.5.3: Typical flow rates vs W:C for a) piston and b) progressing cavity pump
Pumpability of Fresh Cement Paste
The hydraulic properties of grouts profoundly influence the pumpability of
grouts in underground applications. In particular, the transition from a
pseudoplastic to Bingham state causes an abrupt change in pumping rates for many
different types of pumps. While the absolute flow rates illustrated below are
specific to the type of pump, the brand of pump, the power supply (electricity and
voltage, air and P.S.I.), and the attachments (length of hose, etc.), the basic
transition in flow characteristic between the water:cement ratios of 0.35 and 0.4
can clearly be seen and would seem to be independent of the pump used.

∆p
L
Q
D
=
× ×
×
128
4
µ
π
Figure 2.5.4: Theoretical grout pumping performance (efficiency)
Flowability in Grouting Tubes and Breather Tubes
Grout consistency and pump efficiency are often blamed for poor pumping
(grout placement) performance in cases where the real culprit is the undersized
grout tube or breather tube being used. Calling on some basic principles of
hydraulics it is possible to understand why this is the case. Consider a Newtonian
fluid with a viscosity of µ, being pumped through a tube of (inside) diameter, D,
at a flow rate of Q (e.g. litres/sec). The pressure drop per unit length of pipe,
p/L, or the minimum increase in pump pressure required to deliver grout through
an additional unit length of tubing is given by (Streeter and Wylie, 1979):
Note that pressure drop is a function of the inverse of diameter, D, to the power
of 4. This suggests an enormous influence of tube diameter on pumping efficiency.
This relationship is valid only for Newtonian fluids and for laminar (non-
turbulent) flow. A grout of W:C=0.45 can be assumed to fit this description with
minimal error. Figure 2.5.4 illustrates this theoretical relationship, using viscosity
values back-calculated from pumping data by Goris (1990). A grout of W:C=0.4
is also analyzed in a similar fashion although the results are less valid due to
increased grout plasticity. The graphs do, however, reliably illustrate that:
+ Pumping rate decreases and/or pressure drop increases with decreasing W:C.
+ A greater pump pressure is required to pump grout at a higher flow rate. At
these higher pressures, a stronger tube may be required to avoid rupture.
+ The tube diameter has the greatest influence on pumping capability.
The most effective means of improving overall pumping performance is to
increase the breather tube or grout tube diameter (up to a practical
maximum of 20 to 25mm to avoid interference with cable-grout interface).
Constrictions and Flow
Figure 2.5.4 on the previous page assumes that the flow channel (the tube) is
free of constrictions and obstructions. In a real pumping setup and in the borehole
itself, these constrictions and obstacles are common. For example, significant
pressure drops occur whenever the conduit changes diameter. Unavoidable
changes in diameter occur when the grout leaves the grout tube (into the borehole)
and when it enters the breather tube. The latter can be severe. These pressure drops
can be reduced by cutting the ends of both tubes at an angle instead of across the
diameter allowing a smoother flow transition. The number of constrictions and
diameter changes at the pump end should be avoided by matching the pump
outflow and fittings to the grout tube being used. Where this is not possible, a
tapered reducer should be used to reduce the abruptness of the change. Avoid rigid
elbows in the outflow plumbing as well.
Flow in the Borehole
Thin grouts W:C 0.45 possess near-Newtonian qualities (water-like) and can
flow into fractures causing excess grout usage, and can potentially lead to grout
voids, incomplete encapsulation of the cablebolt, and installation hazards due to
roof jacking in laminated ground. Empty cablebolt holes have been observed in
cases where thin grouts were used in highly fractured ground. Semi-thin grouts
0.38 W:C 0.45 may not remain in an unplugged uphole containing plain strand
cables (Toe-to-collar, grout tube installation) although grouts of W:C = 0.4 have
been observed to remain in unplugged up-holes containing modified (nutcaged)
strand. Initial trials may be required before implementing this option.
Flowing grout will take on excess water in a wet borehole resulting in an
undesirable increase in W:C. This mixing action is likely to be more severe in
collar-to-toe installations. Use thicker grouts in wet conditions to compensate.
Very thick grouts W:C 0.30 are unlikely to penetrate and completely surround
the flutes between the wires of the plain strand. In addition, air pockets will be
incorporated into the grout column. This problem may be exaggerated when
double cables are used. Thick grouts will also be unable to penetrate the cages
of modified strand and will form voids around spacers and other fixtures. Semi-
thick grouts 0.30 W:C 0.38 may create difficulties in obtaining return flow
through small breather tubes as discussed and may not completely penetrate
modified strand geometries such as birdcaged and bulbed strand.
0.35 W:C represents an optimum grout for single plain strand cables using a
toe-to-collar, grout tube installation. Encapsulation problems have been observed
when using this grout with double strand cables. 0.4 represents an upper bound for
grout W:C and provides an adequately strong grout with good flow properties for
use in breather-tube installations and where modified strands are to be installed.

L
E L
B F
design
mbedment required
leeding actor
=
−1
Bleeding of Grout
Bleeding of the grout occurs when the particles of cement settle down to the
bottom and water flows up to the top of the grout column. The amount of
bleeding increases with increasing water:cement ratio.
In cablebolt installations, the cablebolt acts as a wick, allowing the water to
flow along the king wire, meanwhile preventing the cement particles from entering
the space between the individual steel wires. The lack of a continuous, protected
king wire along the length of modified geometry cablebolts should prevent some
of the bleeding and grout settlement observed in tests on plain strand cablebolts.
The section of the cablebolt strand which is within the water filled upper section
of the borehole will have no load carrying capacity, and the effective embedded
length of the cablebolt will be reduced from the design length. Excessive bleeding
can disrupt the integrity of the grout column.
Table 2.5.2: Grout bleeding measured in laboratory tests by Goris (1990)
water:cement ratio Bleeding factor (m/m)
0.3 0.016
0.35 0.033
0.4 0.063
0.45 0.096
Results of bleeding tests conducted with both single and twin plain
strand cablebolts; bottom and sides of the cylinder impermeable.
Bleeding factor: Loss of embedment per metre of installed cablebolt.
A limited number of laboratory tests (using a typical paddle mixer and piston
pump) conducted by the authors indicated that when both the upper and lower
ends of plain strand cablebolts drained water from the grout mixture, the amount
of settlement of the cement particles within the column was approximately double
that found in tests where only the top of the cablebolt was free draining. The effect
of lateral bleeding into a fractured rockmass was not considered.
The reduction of the embedment length of the cablebolt due to grout bleeding
can be fairly significant, depending upon the water:cement ratio of the grout. For
most cablebolt orientations, the bleeding of the grout reduces the length of the
grout column at the non-working end of the cablebolt. In this case the effect of
bleeding can be compensated for by increasing the length of the borehole. Use the
following equation to approximate the increase in the hole length required.
2.5.5 Properties of Hydrated Portland Cement
In the cablebolt system, the grout serves to transfer load between the rock and
the cable. The most important properties of fully hydrated (cured) cement paste
which allow it to carry out this function are strength and stiffness. These properties
are functions of water:cement ratio (W:C), cement composition, and elapsed time
since placement.
Compressive Strength
The most convenient index for
describing the strength of cement paste is
its uniaxial compressive strength. A
cylinder or cube of grout is loaded in a
testing machine (A.S.T.M., 1984). The
highest load achieved before destruction of
the sample is the ultimate compressive
strength. It is important to note here that
this value depends to some degree on the
testing setup and on the sample geometry.
Cubes will give somewhat different results
than cylinders and larger samples will give
slightly different results than smaller
samples. It is important to test in a
consistent manner when making
comparisons. The results illustrated in this
section are obtained from cylinders with a
height/diameter ratio of approximately 2.5.
The compressive strength of cement grouts is a function of time, composition,
and of water:cement ratio. The dominant factor in determining the ultimate
strength of hardened (hydrated) cement paste is the initial water:cement ratio
(W:C) of the fresh cement paste. Figure 2.5.5 illustrates an ideal range of
compressive strengths for different grouts prepared in the laboratory. However,
the compressive strength of grout in a borehole in the field may be reduced up to
40% due to inadequate mixing caused by the limitations of the mixer and pump.
There is a clear and approximately linear trend towards increasing compressive
strength with decreasing W:C. Note the increase in variability (scatter) below
W:C=0.38. This is caused by a change in the hydraulic character of the grout
which decreases the mixing efficiency. The reduced water content also means that
less excess water is available for complete saturation and hydration of the cement
particles. Cements at higher water:cement ratios incorporate unused water into
their matrix after hydration, resulting in an increase in micro-voids and reduced
strength.

Figure 2.5.5: Uniaxial compressive strength of grout with respect to water:cement ratio
(after Hyett et al., 1992)
It can be seen from Figure 2.5.5 that the range of W:C = 0.35 to 0.4 provides
the optimum balance of strength and minimized variability.
In addition to the cement composition, the fineness of grind will affect the rate
of hydration and therefore the rate of strength gain. The finer the grind, the more
rapid the strength gain. High-early cements are typically of finer grind but
generally result, however, in slightly lower long term strengths.
Mixing efficiency, chemical variability between cement brands, humidity
during hydration and water quality will also influence cement strength.
Tensile Strength of Cement Grouts
Tensile strength of the grout is defined as the resistance to tensile stress or the
resistance to being pulled apart. This parameter is of minor importance to the
overall performance of the grout in cablebolting applications. There is a great deal
of variability in even the most controlled laboratory testing (Hyett et al., 1992).
The average tensile strength of cement grout of W:C=0.4 is approximately 4 MPa.
There is a slight trend toward higher tensile strengths at lower W:C, but the
inherent variability makes it difficult to specify a quantitative relationship.
Figure 2.5.6: Elastic (Young's) Modulus of cement grout with respect to water:cement
ratio (after Hyett et al., 1992)
Elastic Stiffness (Young's Modulus) of Grout
Along with compressive strength, the elastic stiffness of the grout is one of the
most important measurable grout parameters affecting cablebolt performance.
Young's Modulus is a measure of elastic stiffness and is obtained from the
slope of the graph of axial stress versus axial strain produced during a uniaxial
compression test of the specimen. For the
data shown in Figure 2.5.6, the slope is
measured between points on the curve at
30% and 60% of ultimate compressive
strength. The data shows a clear
relationship which for practical purposes
can be described as linear, with modulus
decreasing as water:cement ratio, W:C,
increases.
While thicker grouts give consistently
higher moduli, the range W:C=0.35 to 0.4
again provides adequate results for
cablebolting.

Figure 2.5.7: Increase in strength and stiffness with curing time for cement grouts
Effect of Curing Time
Normal Portland cement normally takes approximately 28 days (after mixing
and placement to reach complete hydration (effectively 100%) and to obtain
optimum strength and stiffness properties. In underground cablebolting, however,
it is often necessary to bring the bolts (and therefore the grout) into service before
this time. Testing of normal Portland cement (Figures 2.5.7 and 2.5.8) indicates
the following trends:
+ 40-70% strength gain after 3 days. Cables should never be brought into service
(mined through) before this time although plating is possible after 24-48 hours.
+ 80% strength gain after 7 days but with only 50% gain in stiffness. Mining at
this time is not recommended but is possible when necessary due to scheduling
constraints. The use of modified strand reduces the stiffness dependency and
may permit faster cycle times when necessary.
+ 100% strength and stiffness gain after 28 days. Unless otherwise noted, quoted
values for cement strength normally refer to 28 day results.
Higher strength cements such as those containing silica fume can be used to
obtain higher strengths (close to maximum) in 2-4 days. Test results involving
cablebolts are limited in number but such grouts show promise (Hassani et al.,
1992) for accelerating the cablebolting cycle time.
Figure 2.5.8: Increase in cablebolt performance (pullout) with curing time
Effects of Curing time (cont.)
Environmental Sensitivity of Cement Grouts
Cement grouts are susceptible to attack in highly acidic environments; pH6
is considered aggressive for long term exposure (Littlejohn, 1993) although this
is overconservative for mining purposes - most mine waters would fit this
description. Sulphate resistant grouts have a high alkali content and can be used
in these environments (Gunasekera, 1992). Sulphates in groundwater also react
with the tricalcium aluminate in cement to form salts within the cement structure
causing swelling and disintegration. Waters with 0.5 g/litre of sodium sulphate
or 0.25 g/litre of magnesium sulphate are considered aggressive (Littlejohn and
Bruce, 1975). Sulphate resistant cements or blast furnace slag cements provide
some protection. If a problem is suspected, soak grout cylinders in samples of
mine water for time periods corresponding to the expected service life. Check or
test the samples to assess the impact of the corrosive elements in the water.

Blast Damage to Cement Grout
Little testing has been done to evaluate the potential for grout damage due to
nearby blasting (say, in a production stope hangingwall). It is the opinion of the
authors that serious disintegration of cured grouts ( 3 days) occurs only in very
soft or highly fractured rocks. The grout is normally less stiff then the surrounding
rock and as such will suffer less damage as a blast wave propagates through the
rock. Relaxation tests of tensioned grouted cables in an open pit bench indicated
no substantial reduction in cable bond as blasting encroached upon the test bench
(Windsor, C.R. and Thompson, A.G., 1993, personal communication).
When it is necessary to blast nearby before the grout has cured for 3 days,
Heilig and Espley (1993) conclude, based on a study of concrete damage
guidelines and typical blast vibration data from mine sites around Sudbury, that:
Any production blasting within a 24 hour period following grouting of
the last cablebolt should not induce a level of vibration (in any direction)
exceeding 200mm/s at the cable/grout interface......Within 24 hours
following grouting, the minimum distance, in metres, required between
production blasting and any newly installed cables is equal to 5 times the
square root of the charge weight per delay expressed in kilograms and
that for charge weights up to 750kg of explosive per delay; Any open
stope production blasting within 100m of newly installed cables should
not occur within a 24 hour period following grouting of the last cable
and that no production blasting of any weight should occur within 30m
of newly grouted cablebolts for the same 24 hour period.
Oriard and Coulson (1980) investigated the effects of vibrations on mass
concrete and determined that vibrations below 100mm/s between 4 and 24 hours
(minimum of 6 to 8 hours recommended) after concrete placement did not cause
any deterioration. This recommendation is further supported by Esteves (1978)
and Dowding (1985) although damage to the cable/grout bond is not considered.
This guideline can be applied to cables near development rounds. The authors
suggest a 24 hour limit for any larger blasts.
A good rule of thumb is that if the rock suffers blast damage (extensive induced
fracturing), so will the grout. Minimization of blast damage is normally a priority
goal in open stope and drift blasting. Modified (caged or bulbed) strand may
reduce the influence of blast damage on cable performance.
Grout may be damaged, however, if the free end of the cable (exposed at the
collar) suffers severe impact damage. This can occur in a cablebolted back during
a crown blast, if drawpoint cables extend up into the drawcone, in ore passes, or
in drift and sill walls where mobile equipment may impact the free end. The
vibration can be transmitted down the steel and disrupt the cable-grout interface.
Figure 2.5.9: Summary of grout properties for cablebolting
2.5.6 Cement Grout Specifications for Cablebolting

2.5.7 Grout Admixtures
Grout admixtures are normally organic or inorganic chemical substances added
to the grout mixture in small amounts (not exceeding 5% by mass of cement
content) in order to physically alter the properties and behaviour of the cement
paste. In most cases the additive alters the surface properties of the cement
particles during the initial hydration process, then returning them to normal before
final set. In cablebolting applications, admixtures can be used to:
+ Improve pumpability
+ Reduce segregation and bleeding
+ Retard or accelerate the initial set
+ Accelerate the rate of strength development
+ Increase the ultimate strength
+ Induce thixotropic behaviour
+ Inhibit corrosion (grout and /or steel)
While there are hundreds of unique brand name additives currently on the
market, these admixtures fall into the following major categories:
Plasticizers/Fluidifiers/Water-Reducing Admixtures
This class of admixtures allows the use of lower water:cement ratios (and hence
higher strength and stiffness) without affecting mixing, pumping and flow
characteristics. Conversely, these products can create a more pumpable grout
without changing the W:C ratio. It is important to consider all workability
requirements (including the retention of grout in the borehole) to ensure that the
action of these admixtures to solve pumping limitations will not cause unforeseen
problems in other aspects of installation and support performance. The effect of
these products is usually minor, although excessive dosage can lead to variable
setting times, excessive air entrainment and resultant loss of strength.
Superplasticizers/High Range Water Reducers
These admixtures perform the same function as plasticizing and water reducing
agents, but they have a much more exaggerated influence on the hydraulic
properties. These products should be used with care in cablebolt applications since
improper use can have extreme and possibly undesirable consequences. Excessive
dosage can lead to increased variability in setting time, excessive air entrainment,
strength loss, severe segregation and unacceptable retardation of set and cure.
These admixtures should not be used when installing cables in fractured ground
and require tight collar plugging in uphole installations.
Air Entrainment Admixtures
Many admixtures perform this function in addition to other functions listed
here. In concrete construction, air entrainment is used to generate billions of tiny
bubbles in the hardened cement which subsequently act as pressure relief valves
for freeze thaw action. Air entrainment also reduces segregation and bleeding and
improves initial workability. There is reason to expect, however, that air
entrainment may be detrimental to cable/grout interface strength when regular
strand cables are used. The interface mechanics are sensitive to the microscopic
integrity of the cement grout at the cable surface and air entrainment may be
undesirable on this scale. Pull tests should be carried out before implementation.
Retarders
These admixtures are used to delay or retard the rate of set. These may be
useful when pumping many holes from the same batch or where other installation
requirements may cause long pumping cycles. High temperatures ( 30C) can lead
to premature hardening of cement, leading to pumping delays, machinery
breakdown and other problems. Retarders can improve the efficiency of the
installation process in these environments. It is important to obtain detailed
information on these (or any admixture) products. Some retarders may also behave
as water reducers or air entrainers and some may reduce initial (3 to 7) day grout
strength. It is important that ultimate strength is not affected. Pull tests may be
necessary before proceeding. Excessive use of any retarder can lead to extremely
long setting times and unacceptable bleeding or grout flow into fractured ground.
Accelerators
Accelerators are used to increase the rate of initial set and subsequent strength
development. The use of accelerators is not advised in cablebolting applications.
The most common accelerators contain chlorides and nitrates. Chlorides in
particular (calcium chloride is a common accelerator) are extremely detrimental
to reinforcement steel and of course, to cablebolts. One exception is the
accelerator calcium nitrite which also acts as a corrosion inhibitor for encapsulated
metallic elements. Overdose of accelerating admixtures can result in erratic setting
and strength development, cracking and loss in ultimate strength. If acceleration
of set is desired (in cold environments, tight operational schedules, need to plate
immediately) it is more desirable to use lower water cement ratios, or high early
(strength) cement.
Water Retention Admixtures
These reduce bleeding and water loss and improve the uniformity of hydration
and strength.

Thixotropic Agents
These admixtures create a higher W:C cement which behave plastically at rest
and at low pumping rates. Once in motion for a period of time, as in the mixing
and pumping process, the grout behaves in a fluid manner. After cessation of
pumping, or if grout penetrates into a crack (with subsequent decrease in flow
velocity) the grout will return in a short time to a plastic state. This type of
admixture may be desirable for top down grout tube installations with no collar
plug or in highly fractured ground. Continuous mixing of the grout is important
throughout the pumping process. The grout may achieve a false set around the
pump intake or auger feed if mixing is not maintained.
Strength Enhancing Agents
These admixtures can be employed to achieve higher strengths in the cement.
Agents such as silica fume increase ultimate strength in cement, improve
cable/grout bond (adhesional and dilational) and can improve corrosion resistance
of both the cement and the steel cable. There is little experience with this type of
additive in cablebolting but its use is widespread in steel-reinforced concrete
applications.
Corrosion Inhibitors
It is important to define the type of anticipated corrosion (Section 2.4.5):
+ conventional (oxygen and electrolytic) corrosion of the cable steel,
+ acid attack on both steel and cement, or
+ sulphate attack on the cement.
There are a number of additives which can improve resistance to one of these
forms of attack but which may be useless against another. Examples are calcium
nitrite (an active corrosion inhibitor which acts to stabilize oxide films on steel),
silica fume or any additive which reduces bleeding, shrinkage or air entrainment.
Conversely, some admixtures such as accelerators may contain compounds
(chlorides, etc.) which accelerate corrosion and should be avoided.
Expansion Agents
These agents create an expansion in the grout during set and can compensate
for shrinkage and bleeding. They may also increase immediate grout pressures,
improving bond stiffness and frictional resistance to pullout. While some
expanding cements may be appropriate for cablebolting, avoid admixtures which
function by increasing the concentration of micro-voids which could impair the
cable-grout interface strength. These grouts should be evaluated in confined
pullout tests prior to use underground.
General Guidelines for Admixture Usage
Due to the literally hundreds of cement admixtures and brand names available
on the market it is impossible to give guidelines for any particular product. The
following are some general guidelines for usage.
+ It is important to define the desired properties to the supplier. One option is
to determine flow, workability and strength properties in terms of apparent
water:cement ratio. For example, you may specify that you wish to pump at a
W:C=0.45 efficiency, achieve the uphole stability of a 0.35 cement, while
maintaining the long term strength of a 0.3 cement.
+ Check into the experience of the admixture supplier with underground
environments and with reinforcement grouting. If this is limited, the admixture
must at least be approved for application in steel reinforced concrete works.
+ Always perform complete testing for an untried admixture (pumpability,
grout flow in tubes and in boreholes, grout strength and pullout resistance).
+ Obtain complete specifications for the admixture and obtain complete usage
instructions, including mixing sequence, dosages and safety considerations.
+ Many admixtures are originally developed for uses beyond cement
modification and suppliers may not have cement-related safety information.
Many admixtures may increase the existing hazards of cement mixing (e.g. skin
irritation) or create new hazards. Be attentive to complaints from crews when
initiating an admixture program and perform a safety investigation beforehand.
+ Most additives tend to negate the effects of others or may create disastrous
consequences in certain combinations. Do not mix different additives and avoid
mixing even if they perform the same function. Most suppliers have their own
formula for an admixture and this may conflict with that of another supplier.
+ Never exceed the recommended dosage and always follow the correct
mixing sequence. Some agents must be added to the mixing water before the
cement while others are added to the cement paste during mixing.
+ Never use any admixture which contains agents which may enhance
corrosion of steel. Calcium chloride (or any other electrolytic salts such as
sulphates, sulphides or nitrates) must be avoided.
+ Many admixtures are organic and have specific storage requirements. Others
may be temperature or light sensitive during long term storage. Powdered
additives should never be allowed to absorb moisture before use. Liquid
admixtures should not be allowed to freeze. Do not use spoiled admixtures.

Figure 2.6.1: Possible cablebolt failure modes (after Jeremic and Delaire, 1983)
2.6 Load Transfer
Load is transferred between two separating zones of rock through tension in the
cable strand. The cable-grout-rock interfaces must also bear this load transfer. Five
modes of grouted cablebolt failure can occur (Figure 2.6.1):
(A) by rupture of the steel tendon
(B) at the cable/grout interface
(C) through the grout column
(D) at the grout rock interface
(E) through the rock surrounding the borehole.
Strand rupture (A) occurs if the shear loads acting over the embedded surface
area of cable exceeds the maximum tensile capacity of the steel strand. Steel
strand capacity has been discussed in Section 2.4.
If strand rupture does not occur, it has been consistently shown (Stillborg,
1984; Yazici and Kaiser, 1992; Hyett et al., 1992; Goris, 1990; Reichert, 1991)
that in hard-rock applications, failure will first occur along the grout/cable
interface (bond failure) due to inadequate shear resistance or bond strength when
modified (Section 2.9) or plain strand cables are used (B). Other modes of failure
(primarily D E) can occur in soft, weak rocks (Franklin, pers. comm; Carter,
1995) when modified strand cables are used to increase the bond strength.
Figure 2.6.2: Embedment length
and cable capacity
2.6.1 Bond Strength
For cable installations in hard rock, cablebolt capacity is defined by the
properties and strength of the steel tendon, and by the resistance to slip along the
cable/grout interface.
The ultimate tensile capacity of a steel tendon under load is a standardized
specification for the manufactured product. The minimum design specifications
for 15.2 mm steel strand cables are 20 tonnes (~ 200 kN) at yield and 25 tonnes
(~250 kN) tonnes at rupture.
The bond strength of a cable is defined as the resistance to slip (at the
cable/grout interface) along a unit length or a unit surface area of cable. It is useful
to think in terms of load/length required to cause slip. Convenient units are kN/m
or the mass equivalent, tonnes/m. While the actual relationship between ultimate
capacity and grouted length is not always linear, the concept of a normalized bond
strength serves to simplify analysis and design.
Embedment length is a term used to
describe the active length of grouted cable
under unidirectional slip. Consider the
simplest example of a slab of cablebolted
rock displacing from the excavation face. It
is apparent that there are two distinct cable
segments to consider. The loading section
includes the grouted length inside the slab
(embedment length = L ). The remaining1
length, L ( L in this example), forms the2 1
anchor section. These two segments or
embedment lengths can be considered
separately to determine the maximum bond
capacity for the system. In the absence of
external influences or surface fixtures, it is
clear that the shorter embedment length in
the slab will dominate the overall behaviour
of the system.
The total pullout load of a cable increases with active embedment length up to
a limit defined by the steel capacity. For example, a cable with a bond strength of
98kN/m or 10 tonnes/m would reach a maximum pullout resistance of
approximately 250 kN (25 tonnes) over an embedment (grouted) length of 2.5 m.
For embedment lengths greater than this, the cable tendon would rupture during
pullout. The minimum embedment length at which cable rupture occurs during
pullout (e.g. L in above example) is called the critical embedment length.C

Figure 2.6.3: Critical Bond Strength, CBS, for cablebolts - gravity loading
Critical Bond Strength
The minimum bond strength required over a unit embedment length to sustain
a given load (for a given density of rock, for example) is called the critical bond
strength. If the actual (calculated in the following sections) or measured bond
strength is less than this minimum, design adjustments will be necessary.
For example, cables spaced at 2 m x 2 m in a horizontal roof in a rockmass with
a specific weight of 30 kN/m require a critical bond strength of 120 kN/m (2 ×3
2 × 30) or approximately 12 tonnes/m. The case of a rock slab can be used to
illustrate the concept of critical bond strength as shown in Figure 2.6.3.
Select the curve corresponding the unit weight of the rockmass and determine
the critical bond strength for a given cablebolt spacing. If the actual normalized
pullout strength (load/length) is less than this value, the rock will slide off the
cable (bond failure) under gravity loading.
2.6.2 Bond Strength of Plain Strand Cablebolts
Adhesion and Bond Strength
In the context of steel reinforcement of rock or concrete materials, adhesion
describes a bonding mechanism (Farmer, 1975; Littlejohn and Bruce, 1975) in
which a pseudo-chemical bond develops at the steel/cement interface which is
brittle (no residual bond after rupture) and independent of confining pressure
(stress normal to the interface).
Typically, for regular carbon steel and cement grouts with W:C in the range of
0.35 - 0.5, this adhesion or shear resistance is equivalent to 1 to 3 MPa. Over the
surface area of a 15.2 mm diameter cable, this is equivalent to a capacity of 10 kN
over a 20 cm length of grouted cable. Unfortunately, this adhesion is exceeded
after less than one fifth of a millimetre of relative slip (Fuller and Cox, 1975;
Hyett et al., 1992; Nosé, 1993). As such, it is unlikely that adhesion can act
simultaneously over any appreciable embedment (grouted) length and rarely
accounts for any significant percentage of the instantaneous pullout resistance
(bond strength). In fact, as the cable is loaded and begins to slip at the cable/grout
interface, a wave of localized adhesion failure propagates down the cable away
from the loading site.
Adhesion is thereby rapidly removed from the system as this initial bond is
broken and is not considered hereafter as a load transfer mechanism.
Slip, dilation, friction and bond strength
The helical, multi-wire nature of the cable surface creates a negative relief of
equivalent geometry in the hardened grout. After adhesion is removed from the
interface, the cable slips with respect to the grout annulus. If rotation of the cable
during pull-out is prevented, a geometric mismatch occurs between the cable flutes
and the corresponding grout ridges. This mismatch increases with increasing
relative slip as illustrated in Figure 2.6.4.
As the grout ridges must ride up and over the cable wires, the grout compresses
in the confined borehole and thus generates a normal pressure on the grout/steel
interface.
Friction (pressure dependent shear strength) thus develops along this interface
providing resistance to further slip. This interaction is called dilation. Dilation is
limited in the extreme by the absolute scale (height) of the grout ridges. In reality,
dilation pressures develop to the point where these ridges crush, reducing the
maximum dilation to less than 0.1 mm for plain strand cable (Diederichs et al.,
1993).

Figure 2.6.4: Dilation and bond strength: modified versus plain strand cable
Dilation is the key to cablebolt performance and is a complex process which is
dependent on grout stiffness, rock stiffness and grout strength. This relationship
will be explored in the next section.
Bond Strength and Load Transfer
Before proceeding with a discussion of bond strength, it is necessary to
understand the process by which load is transferred from the rockmass to the cable
via the shear resistance at the cable-grout interface. As the rock slips with respect
to the cable, shear stresses (load/unit area) are generated at the interface. As these
shear stresses accumulate along the length of the cable due to the addition of
incremental rock loads, the tension in the steel strand increases (for an unplated
cable) from zero at the face to a maximum at some point into the borehole.
Beyond this point (i.e. in the anchor section of the cable) the shear stresses act
in the opposite direction and can be considered as negative. In this region, the
loads accumulated in the bottom portion of the cable are transferred back to the
rockmass and the cable tension drops back to zero at the upper end of the grouted
strand. The following examples illustrate this concept.
Load Transfer Example: Slab or Wedge
In this example, a slab or wedge of thickness, A (less than critical embedment
length), displaces downwards under the influence of gravity. If the ultimate bond
strength along segment A is less than the critical bond strength, the shear stress
acting on the cable-grout interface in section A will become approximately
constant as the slab slides along (and off) the cable. During slip, the tension in the
steel cable rises linearly from zero at the face to a maximum at the separation
plane between A and B. Segment A is called the pick-up length. Note that in the
anchor length, B, the shear stresses act in the opposite direction as the cable tends
to slip down with respect to the rock. Section B, in this example, is long enough
to transfer the load from A back to the rockmass without significant slip (10mm).
The end of the cable in B may or may not displace at all, depending on the length
of B (if A=B, the amount of slip will be equal). The tension in the cable returns
to nil at the top of section B as all of the load is transferred back to the rock.

Load Transfer Example: Fractured Ground
The concept of bond stress and load transfer become slightly more
complicated when dealing with a fractured rockmass, displacing under gravity or
the influence of stress as in this example. Here the displacement profile (Bawden
et al., 1995) of the rockmass is assumed to be non-linear, with maximum
displacement at the face reducing to nil into the rockmass (at the top of section B).
The boundary between the loading section, A, and the anchor section B, becomes
undefined. In the lower section (A) the rock has displaced more than the cable
(with respect to initial conditions). This generates slip and shear loading on the
cable-grout interface and tension in the strand. At some point into the back, the
relative displacement between the cable and the rock is zero. This is the neutral
point (zero shear and maximum tension) and is the boundary between the pick-up
length and the anchor length. Above this point the load is transferred back to the
rockmass (Section B) as the shear reverses direction and the cable tension drops
back to zero.
Figure 2.6.5: Bond model
A Bond Strength Model for Plain Strand Cablebolts
Numerous models have been developed to explain the complex interactions that
occur at the cable/grout interface, and to evaluate the influence of grout quality,
rock properties and changes in stress in the rock surrounding the borehole (Hyett
et al., 1995; Tan et al., 1993; Fuller et al., 1990). The following is a description
of one such model, presented here without proof or mathematical detail to
illustrate cable behaviour. Interested readers are referred to more comprehensive
references (Kaiser et al., 1992; Yazici and Kaiser, 1992; Diederichs et al., 1993)
for details on the formulation and application of this model.
Consider Figure 2.6.5 where the plain
seven-wire strand slips past the grout
interface. The grout ridges ride up and over
the wires, compressing the grout annulus
which in turn pushes against the borehole
wall. This interaction generates a dilation
pressure within the grout and upon the cable-
grout interface. For the purposes of
understanding the model, the cable is
assumed to be a round bar. This bar expands
in cross-section by an amount necessary to
create a dilation pressure, p, equivalent to
that generated by the actual cable. The
expansion of this bar is represented by a
radial displacement, U, at the surface of the
steel (Kaiser et al., 1992).
Thick cylinder equations (Obert and
Duvall, 1967; Popov, 1978) are used to
simulate the combined cable-grout-rock
system. The rock is approximated by a
cylinder of infinite outer radius. In the case
of a pullout test in the lab, the radius of the
confining pipe is used directly.
The model, illustrated in Figure 2.6.6,
incorporates the modulus of the grout and of
the surrounding rock or pipe (the cable is
assumed to be rigid) and relates a
displacement (dilation), U , at the inner1
boundary with an associated increase in
interface pressure, p (slope M=p /U ).1 1 1
This pressure is related to bond strength
through an interface friction angle N.i

Figure 2.6.6: A model for axial pullout strength of plain stand cables (after Kaiser et al.,
1992; Yazici and Kaiser, 1992; Diederichs et al., 1993)
A Bond Strength Model (cont.)
Figure 2.6.7: Interface friction angle
The relationship between radial displacement and interface pressure is
dependent on the elastic properties of the grout and rock and is linear for intact
grout. Normally, a circumferential tension is generated in the grout annulus
resulting in the formation of radial cracks. These cracks in turn reduce the
effective stiffness of the grout annulus and therefore reduce the magnitude of
dilation pressure at a given radial displacement. Once the radial cracks are fully
developed, the relationship again becomes linear but at a lower (less stiff) value
of M. At a limiting dilation, the dilation pressure at the tips of the grout ridges
reaches a level equivalent to the uniaxial compressive strength of the grout causing
crushing of the ridges. This dilation limit is controlled by the grout strength and
is confining pressure dependent. The dilation pressure can be directly related to
pullout resistance by a friction angle, N, for the grout/cable interface. As pressurei
increases, so does the instantaneous pullout resistance or bond strength. In the
model examples, bond strength is represented either as load/length (kN/m) or as
shear load divided by the sample cable surface area (MPa). The bond strength
corresponding to the ultimate dilation limit is called the ultimate bond strength.
The dilation limit has been calibrated using test data (Diederichs et al., 1993)
and can be used in the model to predict bond strength for any combination of key
input parameters. An associated computer program, CABLEBOND (Diederichs
et al., 1992) is used throughout the following sections to illustrate bond behaviour.
Interface Friction and Bond Strength
Laboratory research (Nosé, 1993) and back-analysis (Diederichs et al., 1993)
indicate that the effective friction coefficient (shear resistance / interface pressure)
between the steel strand wire surface and the grout is approximately 0.4,
corresponding to an average friction angle of 21 to 23 degrees. This range has
been independently verified by Hyett et al. (1995).
Light rust increases surface
roughness and may increase this angle
somewhat (Goris, 1990). Heavy rust
or grease, however, may reduce this
angle considerably, seriously
impairing bond strength (LeClair,
1995; Lappalainen and Pulkkinen,
1982). Note that heavy rust which fills
the cable flutes also reduces the
potential for dilation and so has a
compounded detrimental effect on
bond strength. Paint sprayed on the
cable serves as a debonding agent via
the same mechanisms (Windsor and
Thompson, pers. communication).

Figure 2.6.8: Cable strand rotation, embedment length, and bond strength
Cable Rotation and Bond Strength (Response Range)
Maximum dilation occurs when the cablebolt is rotationally constrained during
pullout. When rotation is permitted, the strand will tend to take the path of least
resistance as it slips past the grout interface. Rather than pushing the grout ridges
up and out of the way, the strand will tend to corkscrew out of the grout column.
This results in reduced dilation, interface pressure and pullout resistance (Figure
2.6.8.a). Tests which are not constrained will consistently give lower pullout
resistances than constrained tests. The two results do, however, provide an upper
and lower bound to actual cable pullout performance in the field (Figure 2.6.8.b).
In addition, it should be noted that even in tests where one end of the cable is
constrained, the other end which is drawn in will still tend to rotate. Thus, as test
lengths increase, more of the cable within the sample section experiences a
rotational slip. Non-rotational bond strengths based on short test sections (less than
30 cm) are not valid for longer embedment lengths. The true strengths measured
in laboratory tests are bounded by the upper bound (non-rotating) and the lower
bound (rotating) strength limits predicted by the bond model (Figure 2.6.8.c).
Figure 2.6.9: Dilation without pressure, caused by grout shrinkage or strand contraction
Interface Separation and Bond Strength
The maximum dilation (induced radial expansion) is normally less than 0.1 mm
for plain strand cables and for moderate confinement can average 0.02 to 0.04 mm
(Diederichs et al., 1993; Hyett et al.,1995). Even minor interface separation can,
therefore, be significant. Separation can occur due to numerous influences:
Grout shrinkage and cable bond strength
If significant grout shrinkage occurs, then it is possible that the grout may pull
away from the cable before any cable loading occurs. This separation must be
closed before any dilation pressure can be generated. In the model, this can be
viewed as dilation without pressure as shown in Figure 2.6.9. This results in a
reduction in ultimate dilation pressure and consequently in reduced pullout
strength. Shrinkage is a problem when high water:cement ratio cements are used
or if grouting is performed in high temperature, low humidity environments.
Cable strain and bond strength
As a cable is loaded axially it experiences elastic axial tensile strain as well as
an associated reduction in effective diameter. When a cable exceeds its yield
strength (20 tonnes for a 15.2 mm cable) the rate of strain increases. In a heavily
loaded cable, this plastic radial strain can be significant and can cause separation
at the cable/grout interface, reducing the maximum available bond strength.
Inadequate quality of strand and bond strength
If the cablebolt strand is poorly fabricated, there may be inadequate contact
between the outer helical wires and the inner king wire (Bawden et al., 1995). As
the cable is loaded, these gaps will close, reducing the diameter of the cable strand,
resulting in interface separation or reduced interface pressure. This and the
previous two strength reducing mechanisms are modelled in exactly the same way
- by an initial increment of dilation without an associated pressure increase as
shown in Figure 2.6.9.

Figure 2.6.10: Effect of borehole diameter
Borehole Diameter and Bond Strength
Borehole diameter has an effect
on the overall system stiffness. This
effect is, however, relatively minimal
over the range of hole sizes currently
in use for cablebolting. While
smaller boreholes yield slightly
higher bond strengths under ideal
conditions, grouting difficulties arise
at smaller diameters which negate
this effect. The effect on ultimate
bond strength (pullout resistance
after approx. 40 mm of displacement
in this case) is modelled in Figure
2.6.10 for two example combinations
of grout quality and rock stiffness.
Grout Strength and Bond Strength
As dilation of the cable/grout interface progresses during axial slip, the dilation
pressure increases. As the grout ridges ride over the cable wires, the interface
stresses become focussed within a decreasing contact area. Eventually, the grout
ridges crush and further dilation is prevented. This point marks the theoretical
limit of bond strength. This limit is stiffness dependent and can be expressed as
a dilation limit curve in the bond strength model. The shape of this line is back-
calculated from analysis of over 140 test results (Diederichs et al., 1993).
Figure 2.6.11 shows different dilation limits for different grout strengths. These
strengths are related to grout water/cement ratio as shown in Figure 2.5.5. Clearly,
increased strength results in increased maximum dilation pressure which in turn
yields greater bond strength. Note, however, the practical difficulties (Section
2.5.4 and 2.5.6) inherent in the placement of thick grouts (W:C 0.35).
Grout Stiffness and Bond Strength
Water:cement ratio also controls grout stiffness (Figure 2.5.6), which in turn
affects the radial stiffness of the system (Slope M in Figure 2.6.6). Stiffer grouts
lead to an increase in dilation pressure for a given radial displacement. This leads
to an increase in ultimate bond strength as shown in Figure 2.6.11.
Note that the example pullout response curves in Figure 2.6.11 are for a
specific borehole stiffness (equivalent to a moderately stiff limestone with
modulus E = 13 GPa) and that actual response will be dependent on the rockrock
modulus (or pipe stiffness in the lab) as described in the next section.
Figure 2.6.11: Influence of grout strength and stiffness as determined by water:cement
ratio (after Diederichs et al., 1993)
Grout Quality (Water:Cement Ratio) and Bond Strength

Figure 2.6.12: Influence of rock modulus (borehole stiffness) on system stiffness,
interface dilation and bond strength (after Diederichs et al., 1993)
Borehole Stiffness and Bond Strength
The overall radial stiffness of the system is defined by both the grout stiffness
and the rock stiffness. The slope M of the model decreases with decreasing rock
stiffness as shown in Figure 2.6.12. It should be noted that rock stiffness has a
dramatic influence on bond strength when the modulus of the rock surrounding the
borehole is close to or less than the modulus of the grout. In very stiff rocks, the
grout modulus and strength are the critical parameters determining bond strength.
It is the stiffness of the borehole rock which is important to consider. Joints and
fractures around the borehole can influence this stiffness. If the average fracture
spacing is more than 5 times the borehole diameter or if the rockmass is
moderately stressed, then it can be assumed that the intact rock modulus dominates
the cable behaviour. For higher fracture densities or in low stress environments,
it may be appropriate to use the rockmass modulus estimated from rockmass
classification schemes. When the intact rock modulus is to be used, it is prudent
to use 50-70% of the laboratory stiffness to account for borehole damage.
Rock stiffness can change during the service life of a cablebolt. As the
rockmass is overstressed, creating more fractures or as existing fractures open, the
effective rock stiffness can decrease, causing a drop in cable bond strength. This
effect has been observed in the field (Hyett et al., 1992; MacSporran et al., 1992).
2
1
2
1 1 2
2 2
2 2
E
d
E d d
d d d
R
R BH
P O I
I P P I O( )
( )
( ){( ) }+
=
−
+ − +ν ν ν
Figure 2.6.13: Ultimate bond strength as a function of grout quality and rock modulus;
Note that actual system capacity may be limited by strand tensile strength
Rock Stiffness, Grout Quality and Bond Strength
Hyett et al. (1992) give the following relationship between the borehole
parameters and the specifications for a laboratory pipe test (pullout):
where:
E = Rock modulus E = Test pipe material modulusR P
= Rock Poisson's Ratio = Test pipe material Poisson's RatioR P
d = Borehole diameter d = Inside diameter of PipeBH I
d = Outside diameter of Pipeo
In Figure 2.6.13, ultimate bond strength is taken as bond strength
(load/embedment length) at 40 mm of axial slip. Compare with Figure 2.6.3.

Figure 2.6.14: Comparison of model predictions (after Diederichs et al., 1993) with
selected laboratory and field tests (after Hyett et al., 1992)
The model predictions in Figure 2.6.13 can be used with considerable
confidence due to the fact that they are the result of a calibration process
(Diederichs et al., 1993) incorporating over 140 pull test results spanning a wide
range of key parameters (grout, confining medium, hole size, etc.).
As shown below, the upper and lower bounds (non-rotating and rotating,
respectively) accurately reflect the range of performance encountered in testing
and in field loading situations. The reader is referred to Hyett et al. (1992) and
Goris (1990) for a comprehensive suite of laboratory and field test results.
Figure 2.6.15: Stress change and
interface pressure
Stress Change and Bond Strength
Recent research (Kaiser et al., 1992; Diederichs et al., 1993; Maloney et al.,
1992, Pieterse, 1993; MacSporran et al., 1992; Bawden, 1994; Hyett et al., 1995)
has shown conclusively that stress change in the surrounding rockmass after the
installation of a cablebolt can profoundly affect the bond strength of the cable. In
short, stress increases cause an increase in bond strength while stress decreases can
reduce the strength. In the latter case, it is possible in an initially stressed soft
rockmass, that cable bond strength can be reduced to nil. Many cablebolt failures
observed by the authors (Kaiser et al., 1992; Hutchinson and Diederichs, 1993)
can be attributed to stress decrease across the installed cables.
To understand this mechanism, it is
necessary to consider the sequence of cable
installation. First a borehole is drilled in
stressed rock. The borehole deforms
inward as it is drilled. After creation, the
borehole wall is radially unstressed. The
cable is then inserted and grouted. The
grout cylinder at this time is also unstressed
but is in full contact with the cable and the
rock (it is assumed that the cable tendon is
of standard quality and there is no grout
shrinkage).
During the service life of the cable in a
mining environment, new excavations are
created in the vicinity of the cable causing
a change in the local stress field. The
borehole responds with additional radial
displacement, in general, contracting under
increased stress or expanding under stress
reduction. This time the grout is also
influenced by these deformations and in
turn, the conditions at the cable/grout
interface are altered.
Returning to the model, a stress increase in the rockmass causes a contraction
of the borehole and a compression of the grout. If the cable is assumed to be
comparatively rigid, this effect is modelled as an increase in cable/grout interface
pressure without any dilation as in Figure 2.6.16 a). The result is an increase in
maximum interface pressure (after dilation) for a given system stiffness and
therefore an increase in ultimate bond strength.

Figure 2.6.16: Conceptual influence of stress increase (a) and stress decrease ( b) on
bond strength of grouted plain strand cables (after Kaiser et al., 1992)
Figure 2.6.17: Example of the influence of stress change on predicted pullout load
W:C = 0.4, Rock Modulus = 13 GPa (after Diederichs et al., 1993)
A decrease in stress in the surrounding rockmass results in an expansion of the
borehole as the rock relaxes. As a result, the unstressed grout becomes separated
from the borehole and/or from the cable. This separation must be closed before
dilation pressure can be generated. This effect is modelled by a dilation without
pressure increase as shown in Figure 2.6.16 b). If dilation pressure has been
generated through previous cable slip, rock relaxation (stress decrease) will result
in an instantaneous reduction in the interface pressure and a reduction in bond
strength. The effect of stress change on an example borehole configuration is
modelled by CABLEBOND (Diederichs et al., 1992) in Figure 2.6.17.
Figure 2.6.18: Influence of modulus ratio on
stress change @ interface
Figure 2.6.19: Effect of stress change and borehole confinement (rock modulus) on bond
strength for a grout of W:C = 0.375.
It should be noted that field research (Maloney et al. 1992) and recent
independent work and bond strength modelling (Hyett et al., 1995) has confirmed
this predicted behaviour.
The relationship between stress
change in the rock mass and bond
strength is also dependent on the
relative stiffnesses of the rock and
the grout as shown in Figure 2.6.18
at right. The grout modulus can be
obtained from the water:cement ratio
using Figure 2.5.6.
Example relationships for ultimate
bond strength (lower bound bond
strength after 20-40 mm of slip) for
different rock moduli are given in
Figure 2.6.19.
In fractured rockmasses the rock modulus can be stress dependent. In general,
rock stiffness will tend to decrease with decreasing stress. Softer rocks are more
sensitive to stress change. The combined result of rock relaxation, therefore will
be greater (a greater drop in bond strength) than is shown in Figure 2.6.19.

Load Transfer and Stress Change
It has been shown that a modest stress reduction in the surrounding rockmass
can have a significant and detrimental impact on cablebolt bond strength. This
bond strength reduction translates into an overall capacity reduction by reducing
the load transfer capabilities of the cable grout interface. For example, consider
an initially clamped wedge as shown here. If the wedge is pinched out or becomes
dislodged, the stresses in the wedge will drop potentially to zero resulting in a
stress reduction experienced by the cablebolt as shown. If this stress drop reduces
the interface bond strength below the critical bond strength (for the particular case
being considered) the weight of the wedge cannot be transferred to the cable and
carried up to the surrounding rockmass as intended and the wedge will fall. Note
that the initial stresses have been diverted over the top of the wedge, resulting in
a normal stress and bond strength increase in the anchor section.
Stress change and pullout stiffness
In addition to the ultimate bond strength, the initial pullout stiffness of a plain
strand cablebolt is also affected by stress change in the rockmass (Hyett et al.,
1995). A stress decrease in the rock causes a significant reduction in stiffness
during the first 5-20 millimetres of slip as the induced gap between the grout and
the cable is closed. Similarly, the system is made stiffer by an increase in average
stress in the rock. The interface gains immediate frictional strength (without slip
and dilation) due to increased normal pressure as the borehole contracts. These
effects are clearly visible in the example shown in Figure 2.6.17. Note again that
the magnitude of these changes is dependent on the relative stiffnesses of the grout
and the local rock.
Calculation of borehole stress change
Note that the stress change referred to in this section is the average change in
stress, (. +. )/2, acting perpendicular to the borehole, using the stress statemax min
at the time of installation as a datum. The change in stress along the axis, . ,axial
of the borehole has a minor (and counteracting) effect. This effect is ignored in the
simulations presented here. This simplification results in a maximum 3-5% error
in bond strength calculation for hard rock applications.
Examples of Stress Change and Cable Performance
There are many circumstances which would result in a decrease in rock stress
and in cablebolt bond strength reduction. Stress reductions in excess of 40 MPa
have been measured across installed cablebolts (Maloney and Kaiser, 1991;
Maloney et al., 1992). Even in very stiff rocks, such a reduction can be serious.
Since destressed and unconfined (i.e. unclamped) rockmasses are the most
likely to require support, it is particularly alarming that the plain strand cablebolt
is at risk of losing its bond strength and overall capacity at the very time when it
is needed most.
It may be necessary to consider alternative cablebolt geometries such as
birdcaged or bulbed strand which tend to show less sensitivity to stress change.
The influence of stress change can also be reduced through sequencing. By
properly timing the installation of cables it may be possible to reduce the stress
decreases experienced after installation. It is also prudent to attach plates and
surface anchorage to cablebolts where access permits. The plate and anchor make
the pick-up section of the cable comparatively insensitive to stress change.
The examples given in Figures 2.6.20 to 2.6.22 are commonly encountered in
mining. It is important to recognize these situations and to design accordingly.

Figure 2.6.20: Wedge detachment, stress drop and bond strength loss
Stress Change Example - Wedge
Figure 2.6.21: Mine-by stress shadowing and bond strength reduction
Stress Change Example - Hangingwall Stress Shadow

Figure 2.6.22: a) Stress fracturing, stress and stiffness relaxation and bond reduction
b) Creation of re-entrant corners (noses) and stress relaxation
Stress Change Examples - Fracture Zone; Re-entrant Corners
Summary of Remedial Measures - Stress Change
Stress change occurs in every phase of mining. When potentially detrimental
stress reductions are identified, the following options are available:
+ Plate cables (with barrel and wedge anchor). This surface anchorage is not
sensitive to stress change. Ensure that some length of cable (upper end) is
reasonably unaffected by stress reduction. Otherwise pullout may still be a risk.
+ Use modified strand cablebolts (Sections 2.6.3 and 2.9). These flared strand
bolts are much less sensitive to stress change.
+ Adjust sequencing to avoid installing cables in high stress zones (e.g. ahead of
an advancing stope front) which will be subject to future stress reduction.
Figure 2.6.23: Commercially available versions of modified geometry strand (Canada):
a) Birdcaged cable b) Nutcaged cable c) Bulbed strand
2.6.3 Modified Geometry Strand
While the plain strand cablebolt has seen many years of successful application
in civil engineering construction and in mining, the acute sensitivity of the plain
strand to imperfect quality control, stress changes and rock modulus reduction
after placement creates difficulties in mining where these problems are common.
For this reason, various modified geometry cablebolts (modifications of the
plain strand) have been developed over the years (summarized in Windsor, 1992)
which possess reduced sensitivities to these elements and which in general possess
enhanced bond strength and stiffness characteristics. Some of the more recent
developments are detailed in Section 2.9.
In general modified cable strands possess enhanced dilational properties. That
is, they serve to greatly increase the geometric mismatch between the cable and
the grout, generating increased pullout resistance. Shear through the grout takes
a larger part in the overall failure mechanism (Bawden and Hyett, 1994) resulting
in higher bond strength and shorter critical embedment lengths (consistently less
than 0.3 m required to break the strand during pullout).

Figure 2.6.24: Examples of increased bond performance of modified strand (pullout)
Modified Strand - Flared Geometries
While some attempts have been made to improve the interface strength through
the attachment of swaged (pressed) anchors (Schmuck, 1979; Goris, 1990),
internal double-acting barrel and wedge anchors attached to the cable (Matthews
et al., 1983), and through the use of single-acting cable wedges (Gendron et al.,
1992) and internal ferrules similar to the nutcage (Windsor, 1990), the flared
geometries illustrated in Figure 2.6.23 are the most common, commercially
available configurations of modified strand (in North America and Australia).
The birdcage (Hutchins et al., 1990) is formed by unravelling plain strand and
then rewinding the wires slightly out of phase with each other, creating a cage-like
structure which is held together by wire ties. The nutcage (Hyett et al., 1993) is
formed by unwinding plain strand, sliding a nut (or series of nuts) onto the king
wire and then rewinding the strand, preserving as much as possible the original lay
but creating a rigid bulb enclosing the nut. Finally the bulbed strand (Garford Pty.
Ltd., 1990; Bawden and Hyett, 1994) is formed by clamping a section of plain
strand between two hydraulic grips and crimping the intervening section to create
a deformed bulb. This process, if performed correctly does not damage the strand
and has the advantage of preserving the tight wind of the rest of the cable.
Modified strand will usually cost more than plain strand (since the plain strand
is the raw material for their manufacture). Volume production, increased
availability and reduced costs, combined with the increased performance where
bond stiffness and bond strength are critical, have already made these products
competitive alternatives to plain strand.
Load Transfer and Modified Geometry Cablebolts
The flared elements of the modified strand serve as concentrated dilation and
load transfer sites along the cablebolt. It has been shown that a single bulbed,
caged or nutcaged element (node) is capable of generating full tensile capacity in
the strand before bond failure occurs (Bawden and Hyett, 1994). Note that the
modified strands are, in general, considerably stiffer in pullout than plain strand,
generating and transfering loads over much smaller degrees of cable-grout slip.
This property is desirable to reinforce fractured ground and to limit displacements.

Figure 2.6.25: Example situations where debonded strand segments are desirable
2.6.4 Debonding
In highly stressed fractured ground, across mobile shear or
delamination structures or in areas with the risk of
dynamic loading from seismic activity, it is sometimes
desirable to reduce the stiffness of the cablebolt system
over a finite central length of strand while maintaining
bond strength at the ends (Figure 2.6.25).
This is accomplished through the use of debonding.
For plain strand sections this can be accomplished with
varying degrees of efficiency through the use of paint,
grease or plastic tubing. The latter is recommended as
the more predictable method. Figure 2.6.26 shows the
expected elastic and inelastic stretch (relative
displacement) along varying lengths of debond.
Where debonding is used in fractured ground it may
be advisable to plate the exposed end of the cable if
access is permitted. In remotely installed hangingwall
fans, it may be necessary to use specialized cables with
a central portion of debonded plain strand between two
ends of modified geometry or a modified loading
section and plain strand anchor.
Such specialized strand (at right) can now be
manufactured by cable suppliers on special order.
Figure 2.6.26: Supplemental displacement provided by debonded strand sections
Figure 2.6.27: Stiffness reduction and increase in displacement capacity of a birdcage
(B.C.) strand with 3m of debonding between two 1m embedded lengths
Debonded Length
The overall stiffness and therefore the total relative displacement in the cable
will include the response of both of the embedded sections as well as the debonded
length. The increase on displacement capacity (ductility) for a birdcage stand is
shown in the example in Figure 2.6.27. Note that under dynamic loading, plastic
cable strain may localize and reduce the available displacement shown here.

2.6.5 Double and Multiple Strand
The bond strengths and tensile capacities
given in this section generally refer to single
strand cablebolts; that is cablebolts consisting of
a single grouted strand per borehole. It is
common in underground engineering to employ
the use of double or twin strand (two strands per
borehole) and, in open pit and civil anchorage
applications, to install more than two strands in a
cluster. The primary motivation for doing so is
the need to increase the tensile capacity of the
steel strand (single strand capacity × number of
strand per hole) for example, when strand rupture
has been observed. It should not be assumed that
the bond strength (load per unit length of twin
plain strand) will be twice that of single strand.
While double strand cables can increase the
bond strength (load per unit length of double strand cable) up to 100% (Goris,
1990) in laboratory testing, it is unlikely that double strand bolts can compensate
for poor bond strength in the field to such a large degree. The factors which
contribute to observed bond performance problems (grout voids, stress change,
confinement, quality control) of single strand are likely to be exacerbated by the
use of double strand. In particular, the interference between strands can lead to:
- Grout void formation between the strands
during placement (see Section 2.12).
- Grout crushing between cables during pullout
(Hutchinson, 1992), resulting in internal
relaxation and reduced bond strength.
If spacers are employed (at right) at 1 to 2 m
intervals along the double strand, then it is
possible to rely on increased bond strength (up to
100%) in design calculations. Spacers maintain
separation between multiple strands and allow
grout to penetrate between the strands and the
borehole wall. Full encapsulation is required for
optimum bond strength.
Bond strength is unlikely to be a critical
concern when multiple modified strand are
employed with good quality control (Bawden and
Hyett, pers. communication).
Figure 2.6.28: Modified geometry double or twin strand cablebolt configurations
currently in use in underground and open pit mines
Figure 2.6.29: Influence of double strand use on cablebolt system capacity
Double Strand Cablebolts
If it can be assumed that the bond strength is optimized in the multiple strand
configuration, then the tensile capacity and the system stiffness per borehole is
increased by a factor corresponding to the number of strand in the hole (Anderson
and Grebenc, 1995; Villaescusa et al., 1992). Installation, experience and quality
control constraints currently limit underground installations to double strand.
Figure 2.6.29 illustrates the relationship between double strand usage and
cablebolt spacing requirements or maximum system capacity for gravity loading.
In essence, if it can be assumed that:
+ the ground is competent enough to maintain surface integrity between
neighbouring cablebolt holes,
+ the bond strength has been optimized,
+ the loading conditions remain the same;
it is possible reduce the number of cablebolt holes by as much as one half when
double strand cablebolts are employed, significantly reducing installation costs.
Always exercise caution when expanding existing bolt patterns.

τ
α
G ultimate
GUCS
( ) =
T D QG ult= π τ ( )
2.6.6 Grout and Rock Shear Strength
In Figure 2.6.1, a number of possible shear failure modes were introduced for
the cablebolt/rock system. The forgoing discussion has focussed on the cable/grout
interface. For plain strand cablebolts in hard rock, this interface is critical. In
weaker rocks or when modified strand is used, shearing may occur through the
grout itself, at the grout/rock interface or through the surrounding rockmass,
allowing the cablebolt and grout column to pull out of the rock mass. It is
important in these situations to consider the available strength of these interfaces.
Grout Shear
The simplest and most conservative approach for assessing the shear strength
of the grout column is given by Carter (1995) and Littlejohn and Bruce (1975).
The strength per unit area is given as:
where UCS is the grout unconfined compressive strength from Figure 2.5.5 andG
is a factor which Littlejohn and Bruce (1975) specify as 10 for civil engineering
applications. For single strand cablebolts or double strand cablebolts with spacers
in mining applications, the authors of this handbook recommend a value of
ranging from 5 in very stiff (hard) rock to 10 in soft or fractured rock in order to
account for grout confinement and frictional strength.
Normalized grout strength, T (allowable load/unit length of grout column) is;
where Q1 is a quality control factor which is equal to 1 for perfect quality
control (Gerdeen et al., 1977). D is the diameter (in metres) of the relevant shear
interface. For plain strand use the conservative value corresponding to the cable
diameter (0.016m). Use twice the strand diameter for double strand cablebolts.
Without considering quality control during grouting, a plain strand 0.35 W:C grout
(UCS = 60 MPa) in moderately stiff rock (=6) would have a limiting grout
capacity of *(0.016)(60/6)(1)=0.5MN/m. If this is compared to cable/grout bond
strength values from Figure 2.6.13 of 0.3 to 0.4 MN/m, it can be seen that the
cable/grout bond is critical.
For bulbed, birdcage or other modified strand, use the outer diameter of the
deformed element (for example, the diameter of the bulb = 0.025m to 0.035m) as
D in the above equation. For the example above, the grout capacity increases to
0.8 MN/m for the smaller bulb.
( )
τ
α
β
φR ultimate
RUCS
( )
tan
=
+
⎛
⎝
⎜
⎜
⎜
⎞
⎠
⎟
⎟
⎟2 45 2
UCS s UCS s eR L
RMR
= =
−⎛
⎝
⎜
⎞
⎠
⎟
( ) where
105
9
T D R= π τ
Grout/Rock Interface and Rock Shear Strength
After checking for grout shear capacity, check the shear capacity at the
grout/rock interface, using the rock properties (Kenney, 1977; Goodman, 1980):
where UCS is the unconfined compressive strength of the rockmass, variesR
from 1 for fresh rock to 2 for moderately weathered rock to 10 to completely
weathered rock, and $ varies from 0.3 to 0.9 for smooth to roughly drilled holes
respectively. N is the friction angle of the rock taken from 15 for weak clay richo
or schistose rocks to 35 for competent granular or crystalline rocks (Barton,o
1974). UCS is taken as;R
and where UCS is the intact (laboratory) rock strength which can be estimatedL
from Table 2.14.1 and RMR is the 1989 Rock Mass Rating from Section 2.14.4.
This shear strength value is again used in the equation:
to obtain an estimate of the shear capacity of a unit length of grouted borehole
(MN/m) where D is now the diameter of the borehole.
Littlejohn and Bruce (1976) recommend that the values of T calculated above
should be divided by a safety factor of at least 2 to account for the simplifications
and uncertainties involved.
It should be noted that grout/rock or rock shear is seldom the observed failure
mode in hard rock mining applications where single or double strand cablebolts
are used. When a larger number of cables are clustered in a single anchor as in
open pit or civil construction applications, or in especially poor rock conditions,
it is critical that the rock strength and associated pullout resistance be evaluated.
The calculations given here are grossly oversimplified but are normally
adequate for design purposes in mining. The assumption of uniform load transfer
leading to the unit shear capacity concept (MN/m) is a key point here. Many
researchers including Farmer (1975), Aydan and Kawamoto (1992), St. John and
Van Dillen (1983) and others have developed more rigorous approaches to the
evaluation of annular shear behaviour for grouted tendons. Littlejohn and Bruce
(1976) and Barley (1988) have also tabulated a variety of pullout test data.

2.6.7 Load Transfer and Surface Anchorage
Plates can be attached to the exposed ends of cablebolts using a barrel and
wedge anchorage system. It is always beneficial to attach plates where access,
timing and economics permit. This is particularly true for overhead installations
and for plain strand cables. The direct rock-cable connection provided by plates
and anchors reduces the dependence on bond strength. Bond strength near the face
can be reduced by grout voids and due to the ungrouted length at the collar
(breather tube installations). In addition, stress change and modulus reduction as
the rock deforms will reduce bond strength where and when it is needed the most.
If plates are attached, load is generated immediately and if the anchor is designed
with a higher capacity then the strand, full tensile capacity of the cablebolt will be
made available. The cable/plate/anchor system still relies on a functional anchor
length up the hole to complete the load transfer process. For this reason, plates
cannot be used as a substitute for quality control and good design.
2.7 Surface Anchorage and Retention
It is often necessary, and when access permits, usually desirable, to complete
the cable system by attaching plates to the end to provide a positive connection to
the face. Screen or straps can complement the plate to retain small blocks between
the cablebolts or between any intermediate rock bolts or rebars. Note that screen
is usually not effective when anchored at points equivalent to normal cable
spacings (e.g. 2m x 2m) and may require rock bolting between cables.
2.7.1 Plates
Plates can be used simply to attach screen or
straps to the cable system. This is valid where the
integrity of the face is such that little loading will
develop near the exposed end of the cables. In this
case, thinner plates may be used. Thinner plates can
be deformed during manufacture to increase their
retention capabilities or to increase the initial
displacement in high-stress or dynamic conditions.
Where the plate is expected to form an integral
part of the cablebolt load transfer system, as in
highly fractured ground, thicker plates must be used
which have pull-through capacities equivalent to the
tensile strength of the cablebolt (200-250 kN).
Table 2.7.1: Typical dimensions and capacities of bolt
faceplates (after Douglas and Arthur, 1983).
Working Load of Bolt
( kN )
Size of Plate
(Length or Diameter)
( mm )
Thickness
( mm )
80 125 to 150 7
150 150 to 200 10
300 200 to 250 * 12 *
* Recommended for cablebolting in fractured ground
Note that in order to prevent punch-through of the barrel and wedge anchor
fitting, the hole in the plate should be only slightly (1 to 2 mm) larger than the
diameter of the cable strand (15.2 mm). An exception is when the cable is angled
with respect to the face. The edge of the hole must not pinch into the cable and
therefore, must be larger than this limit. A bevelled or cupped washer should then
be used to provide full bearing on the plate and to provide a reduced diameter to
receive the anchor. If rounded barrels are to be used directly on a plate surface, the
plates should have matching spherical insets to receive them.

Figure 2.7.1: a) Typical Barrel and Wedge (BW) anchor assembly for cablebolts
b) Schematic plating/tensioning configuration (after Thompson,1992)
2.7.2 Surface Anchorage - Barrel and Wedge
Cablebolts cannot be threaded to receive a locking nut for surface fixture
placement and anchorage. Some suppliers have marketed cable elements with
threaded end caps, pressed or welded onto the end of the strand, for affixing plates
or straps. The conventional method, however, for surface fixture attachment and
tensioning involves the use of a concentric barrel and wedge assembly which is
activated by tension in the steel cable applied by a specialized jack.
A typical barrel and wedge anchor assembly is shown above. The conical
wedge assembly can be composed of two or three components held together by a
spring wire or rubber ring. The inside surfaces of the wedge are serrated with
concentric rings of teeth to grip the cable. The wedges press against the barrel as
they are driven inward by the jacking unit or pulled in by the relaxation of the
tensioned cable. As long as there is residual load (tension) in the cable, the anchor
is kept in place as the cable continues to pull against the wedges. The outer barrel
completes the load transfer to the plate which bears against the rock.
Barrel and Wedge Anchor Specifications
There are currently no standards for barrel and wedge anchors in mining. For
effective anchor performance, however, a few general manufacturing and
inspection guidelines can be specified:
+ The optimum wedge and barrel length is approximately 2.5 to 3.5 times the
diameter of the cable (i.e. 35 to 55 mm ).
+ The wedges and barrel should be of the same length. If the wedges are inserted
(without a cable) snugly into the barrel, both ends of the wedge cluster should
be flush with the barrel. This will ensure the proper positioning when installed
on a cable. When installed, the narrow end of the wedges should not extend
past the back of the barrel while the wide end should not countersink.
+ The curvature of the wedges should be equal to that of the inside of the barrel
when installed on a cable. This means that without a cable, only the corners of
the wedges touch the barrel. Incorrect curvature leads to wedge failure.
+ The outer diameter of the barrel should be 65-100% greater than the diameter
of the hole in the plate to prevent punch-through.
+ Barrels with rounded backs (on plate side) improve plating performance where
cables are not perpendicular to the rock face. Use rounded barrels only with
matching plates which have a curved recess to accept the barrel. Rounded
barrels on flat plates will result in point loading which can lead to plate failure.
+ Three component wedges improve the efficiency and consistency of grip but
impose additional cost and handling difficulties. Two component wedges are
most commonly used in mining.
+ The outer barrel must be made of a tough steel to resist splitting. There should
be no risk of barrel splitting for cable loads in excess of 25 tonnes.
+ The inside surface of the wedges must have hardened teeth (approximately 1
mm pitch) in order to bite into the cable steel, thereby achieving a positive grip.
The teeth should show no tendency to crush or flatten during installation.
+ The wedge teeth should be 0.4-0.6mm high. The tooth face towards the
borehole should have a slope of approximately 30 degrees with respect to the
cable axis. Away from the borehole the teeth should have a steeper, 60 degree
slope. This ensures efficient jacking and high residual gripping capability.
+ Do not use wedges from one batch with barrels from a different batch or a
different supplier.

Figure 2.7.2: Cablebolt anchorage for: a) Pre- tensioning; b) Cablebolt plating
+ The wedge (and inner barrel taper) angle should be between 7 and 9 degrees
with respect to the cable axis. Most wedges are manufactured at the optimum
angle of 7.5 degrees. Gripping performance degrades with higher wedge angles.
Shallower angles will result in excessive wedge displacement during
installation, require longer barrels and can lead to unpredictable performance.
+ The barrel should provide sufficient thickness of steel adjacent to the wide end
of the wedge - the tapered hole should not extend to the outer edge of the
barrel. This will increase the risk of splitting.
+ The wedge angle must be exactly equal to the barrel taper angle. This can be
checked by inserting the wedges snugly into the barrel (without a cable). There
should be no separation between each of the wedges and between the wedge
corners and either end of the barrel. The wedges will spread when installed on
a cable but the wedge angle will not change. The inside of the wedges should
remain parallel with or without a cable in place.
Installation Considerations - Tensioning and Plating
In civil engineering and in open pit mining applications, the barrel and wedge
anchor is used with a jacking system to generate and to retain a preload in the
cable of 10 to 20 tonnes per 15.2 mm cable. These cables would typically have
ungrouted (or debonded) lengths of 4 to 10 metres between a grouted anchor
section and the borehole collar and plate (the remainder of the hole may or may
not be grouted after tensioning). The purpose of this procedure is to create an
immediate active load at points in an array of cables, in order to directly improve
stability by reducing displacement, enhancing joint interlock and by increasing
normal pressure and thereby friction along potential slip surfaces.
This application differs significantly from the more common underground
mining application - the attachment of plates or straps to the free end of cable
bolts. The load generated in the steel serves primarily to maintain a frictional grip
within the barrel/wedge/cable assembly at the collar. In addition, the active load
on the plate provides limited compression to the immediate surface rockmass,
slightly improving local stability. The cable loads generated in this application
range from 5 to 10 tonnes. A drop in cable tension could result in a loosening of
the barrel/wedge anchor and loss of the plate. The free length of effectively
ungrouted cable includes any length of (non-grouted) packing at the collar and
some unknown length of cable in the grout column adjacent to the collar
(dependent on cable type, cable installation method, grout quality, load magnitude
and time elapsed since grouting). Average effective free lengths can range from
0.5 to 2 metres.
The difference in free length and the magnitudes of the applied and residual
loads are the most significant differences between the two applications and
determines, for the most part, the type of jacking system to be used.
The load retained in the cable is critical for the effective gripping of the barrel
and wedge anchor. If the anchor meets the desired specifications, this post-
installation load is influenced by the jacking configuration, the jacking load, the
effective free length of the cable, the surface character of the rock face and the
presence of screen or soft backing behind the plate. If this load is inadequate, the
anchor may loosen during blasting or seismic activity. Improperly installed
anchors have been observed by the authors to fall off or creep down the cable
under dynamic loading or by static load generated by near-surface rockmass
displacement.
A number of jacking configurations are in current use. The type of jack can
profoundly affect the quality of the anchor installation. Research performed by the
CSIRO (Thompson, 1992) has shed considerable light on this issue. The results
of this work are summarized here with respect to plating applications. Note that
configuration A in Figure 2.7.3 is recommended for underground cablebolt
plating applications.

Figure 2.7.3: Jacking Configurations
(after Thompson 1992)
Jacking Configurations
There are several options for the application of the reaction load between the
jack and the anchor assembly (Pw is the percentage of the total jack load, Pt which
is applied directly to the wedges and represents the main difference between the
four configurations).
A. 100% of the jacking load applied to
the wedges. In this case, the wedge
load, Pw = 100% (of the total jack
load).
Recommended for Plate Installation
B. 100% of the jacking load applied to
the outside of the barrel with no load
applied to the wedges (Pw = 0%).
Highly Undesirable
C. Load applied to the barrel with a
fixed spacing nose cone partially
loading the wedges (0% Pw
100%).
D. Load applied to the barrel with a
spring loaded nose cone applying
partial (and consistent) loading to the
wedges (Pw.10%).
Recommended for Pre-Tensioning
Figure 2.7.4: Theoretical relationships between initial wedge load, Pw (% of total Pt),
peak load, Pmax, during installation and the residual cable load, Pres, for
a nominal jacking load, Pt of 10 tonnes (after Thompson, 1992)
Three key considerations determine the correct reaction configuration:
+ The percentage, Pw, of the total jacking load which is applied directly to the
wedges during the installation process.
+ The relative magnitude of peak cable load, Pmax, generated behind the anchor
(in the borehole) compared with the total load generated by the jack.
+ The residual load, Pres, remaining in the cable after the jack is removed.
When the reaction load is applied directly to the wedges (A: Pw=100%), the
wedges lock into the cable and barrel before the peak load in the jack is reached.
This lock-in prevents additional load transmission to the cable behind the anchor.
Instead, the jack continues to load only the length of cable between the anchor and
the jacking grip. The peak cable load behind the plate will be 30 to 50% of the
peak load registered by the jack even for very short free lengths (Line A in Figure
2.7.4). Nevertheless, when the jack is removed the load in the cable behind the
plate remains relatively constant with little relaxation. This is because there is little
or no additional draw-in of the wedges after jack removal.
Jacking units currently available with spring-loaded nose cones apply a
maximum of 1 tonne (10 kN) onto the wedges while applying a total of 10 tonnes
to the anchor assembly. In this case, when only a partial load is applied to the
wedges (Pw=10%) the peak load in the cable behind the anchor approaches the
maximum total load registered by the jack as shown in Figure 2.7.4 for longer free
lengths. This is a substantial increase in efficiency (Line D).

With this second jack there is, however, the increased probability of additional
draw-in of the wedges after the jack is removed. If there is a long free length (4
to 10 m) this draw-in is accommodated with little relaxation of load in the cable.
If the free length is shorter than 3 metres, this draw-in can result in significant load
relaxation. In addition, the nose cone spring in current models has been observed
to fatigue, causing Pw to drop to 0%, and lose its effectiveness over time and
extended usage (Windsor and Thompson, pers. communication).
It should be noted that jacking systems which apply 100% of the load solely to
the barrel (Pw=0%) are totally unacceptable for plating applications (Line B). The
residual load in the cable can approach zero in this configuration. In tensioning
applications with longer free lengths, the results can be unpredictable.
A nose cone with a calibrated recess to apply delayed loading to the wedge can
be effective for tensioning applications (Line C). When plating cables with short
free lengths, strict dimensional tolerances between the nose cone and the anchor
are required. It is advisable to purchase the jack and anchors from the same
supplier in this case, and to ensure perfect matching between the two.
Recommended procedures
Now that the basic mechanisms involved in the jacking process have been
examined, it is possible to recommend the appropriate jacking system to ensure
that the capacity of the anchorage is maintained or that the desired pre-tension in
the cablebolt is achieved:
Plating and surface fixture attachment (Free length 2 m)
Jacks which apply full load to the wedges during installation are recommended
for barrel and wedge anchorage of surface fixtures where a residual cable load of
4 tonnes is required. Note that using this method, a nominal jacking load of 8 - 12
tonnes (2-3 times the desired residual load) is required due to the inherent
inefficiency of the system. Inadequate loads lead to anchor failure.
Tensioning (Pre- or Post) (Free length 4 m)
Jacks which apply a partial load to the wedges by means of a nose cone spring
are recommended for tensioning applications where high residual loads are
required and most importantly, where long free lengths are present. If the spring
fatigues with use (as is the case in current designs) then the wedge pressure will
approach zero and the consistency of the installation will degrade and may not be
acceptable. The spring should not be compressible by hand - it should exhibit only
a slight give. Maintenance of the spring is especially critical if such a jack must
be used for plating or tensioning of cables with less than 4 metres of free length.
In this case the rigid nose cone (Pw=100%) is preferred.
Key plating and tensioning tips:
+ Never allow oil, dirt or significant rusting to occur on barrels and wedges
before installation. Ensure that the cable segment receiving the anchor is clean.
+ Recommended tension load for plate or strap attachment is 4 tonnes. A rule of
thumb is that 8 -12 tonnes of nominal jack load is required to ensure this
minimum residual cable tension regardless of the installation method used.
+ Do not confuse jack pressure with jack load. Ensure that the gauge is clearly
marked. Ensure that the pressure to load conversion factor (i.e., the area of the
ram in the jack) is clearly affixed or marked on the jack for later reference.
Clearly specify the jacking limit in the same units as the gauge. It is confusing
to specify cable load in tonnes if the jack gauge is in p.s.i. or kPa.
+ Do not exceed yield strength of the cable strand (20 tonnes) during tensioning.
Premature rupture can initiate at the wedges at loads close to yield.
+ When plating, a light bending of plates is desirable as a quality control measure
(post-installation rebound should be minimal). Significant bending (5 degrees)
indicates underdesigned plates or excessive jack loads and should be avoided.
+ When plating against pre-installed screen, jack the cable until the plate touches
the rock. The screen cannot provide the necessary rigidity to ensure anchor
integrity. Blast vibration will shake the anchors loose.
+ When installing plates against rough rock surfaces, it is desirable to apply
sufficient jacking load to crush the tips of sharp surface ridges in order to
establish a more positive contact with the surface. These small rock edges may
otherwise be destroyed by blast vibrations of rock creep, reducing the residual
load in the cable and impairing the load capacity of the anchor. Thick stiff
rubber backings may also be used to dampen blast vibrations where problems
are observed (failed plate anchors) and to allow for limited face displacement.
+ It is useful to affix strong reflective tape or a paint stripe on the cable 5 mm
from the top of the wedges after installation, to serve as a draw-in indicator.
+ Do not attempt to attach BW anchors to modified strand unless a suitable
length of plain, unmodified strand is left at the collar end of the cablebolt
(discuss this with the supplier). Allow enough unmodified length outside the
hole for anchoring and jacking as well as at least 0.5 m inside the borehole.
+ When plating a double strand cablebolt use two anchors if necessary. Avoid the
practice of bending the second strand behind the plate. This leads to an
imperfect contact between the plate and the rock. The preferred option is to
recess the second cable, exposing only a single strand to be anchored.

Figure 2.7.5: Double BW anchor
+ When installing cables over screen, it is preferable that the screen not be cut to
accommodate the cable and plate. Instead, place the plate and anchor over the
screen, mating the retention system to the support system.
+ When screen has been previously stretched flat across a large depression in the
rock surface making tight placement of a plate impossible, the screen should
then be cut to allow placement of the plate and anchor on the rock surface.
Anchors under dynamic loading
Under severe dynamic loading, such as heavy blasting or rock burst (seismic)
conditions, a properly tensioned barrel and wedge anchor with sufficient free
length behind the plate should perform adequately. In the authors' experience,
there have been instances, however, where the anchors lose their grip under severe
vibration. In some cases, the barrel and wedge anchors can be ejected under high
velocity and can constitute a safety hazard (Seldon, 1996, pers. communication).
In these conditions, the first recommended option is the use of domed or curved
plates (Section 2.7.1). The dome must be strong enough to withstand the initial
tensioning/plating load or 10-12 tonnes but should be designed to collapse under
more than 18 to 20 tonnes. This will provide a seismic displacement contingency.
The use of double anchors on single
cables may also provide a belt and
suspenders solution, providing that the
dynamic loading is not enough to break the
cable strand (Figure 2.7.5). The first anchor
will be thrust onto the second anchor
forcing the wedge to grip.The disadvantage
of this system is that it will be difficult to
achieve sufficient residual load P in theres
outer anchor due to the lack of free length.
Another alternative which has not been widely used and which should be tested
before adoption, is the use of a swaged (pressure-fitted) steel anchor (Section
2.9.3) in place of a barrel and wedge. Quality control is difficult to achieve with
this system but the displacement characteristics of such an anchor would be ideal
for plating in dynamic loading conditions. In static load pull tests (Goris, 1990),
the swaged anchors have been shown to yield at a constant load slightly below the
yield point of the strand. This characteristic is ideal for dynamic support. It is not
possible to apply any tension to the plate with a swaged anchor and so the system
will be very soft, applying no initial load to the rock surface.
Steel straps used for support in bursting conditions should have slots through
which the cables are installed. The strap can slip with respect to the cablebolts to
accommodate rock displacement during blasting or seismic disturbance.
Figure 2.8.1: Typical results from direct shear tests of cablebolts (after Windsor and
Thompson, 1993; Windsor, 1992; Windsor et al., 1988)
2.8 Shear Loading of Cablebolts
While most of the research and testing of cablebolt capacity deals with axial
loading (pullout), actual conditions in the field normally include at least some
component of shearing. For this reason, recent research (Windsor, 1992; Windsor
et al., 1988; and Bawden et al., 1994) has focussed on the response of various
cable configurations to different modes of shearing. Section 2.2.2 describes the
testing apparatus used to perform these tests. The following discussion is a brief
summary of cablebolt response and is not intended as a comprehensive reference.
2.8.1 Direct Shear
Windsor (1992), Windsor and Thompson (1993) and Windsor et al. (1988)
describe a testing device for grouted cablebolts subjected to direct shear (Figure
2.2.6). The cables can be oriented with respect to the shearing surface to simulate
pure shear ( noted as 90-), shear + tension (135-) and shear + compression (45-).
Note in Figure 2.8.1 that cables oriented at 135- degrees to the direction of shear,
such that tension and shear are mobilized, give the stiffest results. Cables which
experience compression, kinking and shear (45-) have very soft responses and
reduced capacities.

Figure 2.8.2: Analogous field conditions corresponding to oblique axial/shear testing
2.8.2 Oblique Loading - Shear
Consider the hangingwall block examples shown in Figure 2.8.2. In a
simplified gravity loading scenario, the block moves down under its own weight.
Cablebolts are installed perpendicular to the hangingwall to support the block. The
relative components of shear and axial loading experienced by the cablebolts will
depend on the angle of the hangingwall. For a horizontal surface, the loading will
be purely axial. Shearing of the cable increases with increasing inclination of the
wall and of the separation plane.
Bawden et al. (1994) present preliminary results from an extensive testing
program using the apparatus shown in Figures 2.2.7 and 2.8.2 to investigate this
scenario. A summary of these results is shown in Figures 2.8.2 and 2.8.3. Note
that the ultimate capacities of the strands does not show significant reduction with
increased shear. This is likely due to crushing of the unconfined grout at the
separation plane. For steep angles of loading 45 degrees, the tendency toward
axial pullout is reduced as shearing becomes dominant. In short embedment
lengths this gives the impression of increased capacity.
In shear, the system stiffness over the first 10 to 20
mm of slip is reduced (Bawden et al.,1994). This is a
significant finding and can explain the inability of low
angle cablebolts to effectively reinforce sloughing stope
walls. Such cables are designed to prevent beam
delamination (axial displacement) but are less effective,
for example, when the hangingwall is undercut and
displacements become vertically downward. If
undercutting is suspected, high angle (closer to vertical)
cables should be included in the array as shown at right.
Figure 2.8.3: Example tests results - oblique loading. (after Bawden et al., 1994)
Oblique Loading - Results

Figure 2.8.4: Optimization of cablebolt angle
2.8.3 Cablebolt Orientation
A few simple conclusions pertaining to the optimization of the cablebolt
installation angle can be drawn from the discussions in Sections 2.8.1 and 2.8.2.
+ Cablebolts installed across a confined shearing surface (sustained contact
during shear) perform best when oriented at an acute angle to the direction of
the discontinuity and to the direction of shearing. In this case an angle of 30
to 60 degrees is recommended. Ensure that the orientation is chosen such that
the shearing immediately induces stretch and not buckling in the cable strand.
This will optimize cable behaviour and will also induce confining load on the
shearing and dilating discontinuity, effectively increasing the frictional
resistance. If this orientation is not practical then follow the next point:
+ Cablebolts installed across any surfaces which experience normal separation
(i.e. are pulled apart) should be installed parallel to the direction of
displacement, regardless of the orientation of the discontinuity.
It is therefore important to understand the nature of the displacement across a
discontinuity and to orient the cablebolts accordingly as in Figure 2.8.4.
2.9 Cablebolt Strand Alternatives
The following section summarizes the engineering properties and installation
recommendations for a number of the cablebolt strand options currently available.
Many of the options discussed are readily available from suppliers in North
America, Australia and elsewhere at an increasingly competitive cost. While
15.2mm plain strand cable has formed the basis of much of the discussion in this
book, one purpose of this discussion is to illustrate many of the detrimental
sensitivities inherent in this device. While there have been hundreds of successful
applications of plain strand cable in civil construction and in mining, the diverse
and ever-changing underground environments typical of hard rock mining warrant
the consideration of a wider toolbox of cablebolt options.
Table 2.9.1: Strand summary: (* indicates a detailed summary in the following section)
Plain Strand - 15.2 mm * The basis of all other steel strand alternatives
Modification by Coating Increase corrosion resistance
Epoxy Coated * - Outside of the cable is coated
Epoxy Encapsulated * - Epoxy penetrates internal spaces in strand
Modification by Unwinding Increase Bond Strength Stiffness
Birdcaged Strand * - Disassemble wound plain strand and rewind out of
phase to create an open cage
Birdcaged 14-wire * - As above, but created with two plain strands unwound
and rewound coincidentally
Modification by Inclusion Increase Bond Strength Stiffness
Nutcaged Strand* - Disassemble plain strand; Rewind tightly while
inserting hexagonal nuts at intervals over the king wire
Ferruled Strand - Inclusion is a rounded ferrule instead of a nut
Modification by Deformation Increase Bond Strength Stiffness
Bulbed Strand* - Grip intact plain strand in hydraulic device and
symmetrically kink the wires into a flared bulb
Strand between bulbs is undisturbed
Modification by Attachment Increase Bond Strength and Stiffness (Locally)
Buttons or Swages * - Clamp, press or weld cylindrical steel buttons onto
strand at prescribed intervals
BW anchors - Install pairs of conventional barrel and wedge anchors
such that the wedges of the opposing units press
against each other and lock onto the cable
Modification by Debonding Eliminate bond along partial strand length to increase
displacement capacity
Tubing - Insert tubing over debonding length
Paint, Grease, Coatings - Apply paint or heavy grease to strand prior to grouting
Double or Multiple Cables * All of the above options can be used in tandem (double
strand) or in combination (e.g. one bulbed strand and one
plain strand in each hole). Single modified strands can
include differently modified segments.

2.9.1 Plain Strand
Recommended Grout: 0.35 (Grout Tube) 0.38 (Breather-Tube)
Minimum Hole Single 48 mm 48 mm
Diameter (Sct.2.10): Double 64 mm 64 mm
Recommended Applications:
Moderately blocky ground with limited potential for relaxation after
installation. Semi-ductile interface allows for moderate dynamic
loading. Effective in uphole installation with plates.
Capacity Notes: (references: see Sections 2.4 to 2.6)
Initial Stiffness
- 50 % of Pullout Load after 1 to 2 mm; 80 % after 10 to 20mm
Displacement Capacity
- Dependent on embedment length.
In the event of pullout (no strand rupture), high residual strength
is maintained for 40 to 80 mm. Otherwise strand rupture occurs
within 20mm. Double strand increases stiffness (up to 100 %).
Load Capacity
- Pullout load with respect to embedment length ranges from 20 to
35 tonnes/m (1 tonne ~ 10 kN) for recommended grout range but
is highly sensitive to stiffness and stress change.
- Tensile capacity: Yield = 20 tonnes/ strand,
Rupture = 25 tonnes / strand.
Sensitivity
- Highly sensitive to reduction of rock modulus below 10 GPa.
- Sensitive to stress change - Relaxation of rockmass can reduce
bond strength to near zero if severe.
- Quality control is necessary with respect to grouting and storage.
- Quality and surface condition (clean) of strand critical.
Advantages:
- Readily available, inexpensive and easy to install
- Can be shipped in continuous reel
- Will fit into a smaller hole than modified geometries
- Easily fitted with plates and surface anchors
- Relatively ductile system
Disadvantages:
- Extreme sensitivities as noted
- Lowest bond strength and highest critical embedment length
Figure 2.9.1: Performance summary for plain strand cablebolts
Plain Strand

2.9.2 Epoxy Coated/Encapsulated Strand
Recommended Grout: 0.35 (Grout Tube) 0.4 (Breather-Tube)
Minimum Hole Single 48 mm 48 mm
Diameter (Sct.2.10): Double 64 mm 64 mm
Same as for plain strand.
Provide corrosion protection in aggressive environments.
Epoxy coatings should be certified for the type of corrosion
expected (e.g. acid, electrolytic), for flexibility of coating, abrasion
resistance, etc. Epoxy should have embedded grit.
Epoxy encapsulated cables have epoxy filling the internal voids in
the strand. These channels in coated strand can otherwise
provide concentrated corrosion sites in aggressive environments.
Capacity Notes: (ref: Goris, 1990; Goris et al., 1994; Littlejohn,
1993; Windsor, 1992; Dorsten et al., 1984)
Initial Stiffness
- 70 % of Pullout Load after 1 to 2 mm; 90 % after 10 mm
- Dependent on embedment length.
- Comparable with plain strand although coating may rupture after
large displacements ( 100 mm) permitting concentrated
corrosive attack in aggressive environments.
Load Capacity
- Pullout Load up to 30% higher (with grit coating) than plain strand
- Tensile capacity: Yield = 20 tonnes / strand
Rupture = 25 tonnes / strand
Sensitivity
- Improved corrosion protection.
Advantages:
- Installation procedure same as for plain strand
- Corrosion resistance and long service life
- Slightly improved bond performance
Disadvantages:
- Cannot be plated unless stripped
- Expensive
2.9.3 Swaged/Buttoned Strand
Recommended Grout: 0.4 (Grout-Tube*) 0.4 (Breather-Tube**)
Preferred Method for: *Downholes **Upholes
Minimum Hole Diameter: Allow for button and grout / breather tube
Support across known joint surfaces
Anchorage for debonded cable sections
High relaxation zones
Capacity Notes: (ref: Goris, 1990; Goris et al., 1994;Schmuck,1979)
Initial Stiffness
- Controlled by embedment depth and rigidity of button fixture.
- In the event of pullout (no strand rupture), high residual strength
is maintained for 40 to 100 mm. Otherwise strand ruptures within
10-20mm. Buttons can be designed with yielding limit for ductility.
- Button can be used to provide positive anchorage around a
debonded segment.
Load Capacity
- Pullout increased 150% for button depth 100 mm (Fig. 2.9.3)
- Pullout load reduced from plain strand if button depth 50 mm
- Tensile capacity: Yield = 20 tonnes
Rupture = 25 tonnes.
Sensitivity
- Sensitive to button depth (placement with respect to fractures)
If proximity to fracture 50mm, grout is pushed out of hole
- Sensitive to button fixture quality control
- Sensitive to button spacing and attachment method
Advantages:
- Bond is less sensitive to stress change and confinement
- Provides positive anchorage at button
- Cablebolt can be easily plated
Disadvantages:
- Sensitivity to button placement
- Large borehole required
- Sensitive to voids on lee side of button (w.r.t grout flow)

Figure 2.9.2: Pullout response for epoxy coated strand (after Goris, 1990)
Figure 2.9.3: Pullout response of strand with buttons or swages (after Goris, 1990)
Epoxy Coated Strand
Strand with Buttons
2.9.4 Birdcaged Strand
Recommended Grout 0.37 (Grout-Tube*) 0.4 (Breather-Tube**)
Minimum Hole Single: 64 mm 57 mm
Diameter: Double: 76 mm 76 mm
(Sct. 2.10) 14-wire: 76 mm 70 mm
Recommended Applications
Highly fractured ground
Ground with potential relaxation after installation
Stiff system for immediate load response
Capacity Notes: (ref: Goris, 1990; Goris et al., 1994; Hutchins et
al., 1990; Cortolezzis, 1991)
Initial Stiffness
- Double the initial stiffness of plain strand
- Variable
- Can be high in low W:C grout
Load Capacity
- Pullout load strength in short embedment lengths 35 to 80 %
higher than plain strand.
- Strand ruptures at 20 - 22 tonnes due to eccentricity of loading on
individual wires
Sensitivity
- Birdcages should be firm to the grip. Loose birdcage reduces
effectiveness and increases installation difficulty.
Advantages:
- Only slightly more expensive than plain strand
- Stiff, strong system
- Birdcage has no central channel to contribute to grout bleeding
Disadvantages:
- Cannot be plated unless plain strand section is left at end.
- Difficult to insert and handle
- Cannot be shipped in a continuous coil
- Requires larger borehole
- Response in partial or full shear may be unpredictable due to
uneven loading of wires
- Cannot be installed with standard automatic cablebolt pushers

Figure 2.9.4: Performance summary for birdcaged strand
Birdcaged Strand
2.9.5 Nutcaged Strand
Minimum Hole Single: 51 mm 48 mm
Diameter (Sct.2.10): Double: 57 mm 51 mm
Higher ductility than birdcage and bulbed strand
Capacity Notes: (ref: Hyett et al., 1993; Bawden and Hyett, 1994)
NOTE: Recommended nut size 12 - 16 mm
Initial Stiffness
- Stiffness for embedment lengths of 300 mm are at least 100 %
greater than plain strand. Larger nut gives higher pullout stiffness.
- In moderate confinements at embedment lengths of 300 mm,
cable rupture occurs at 25 - 30 mm for 15.9 mm nut and at 40 to
50 mm for 12.7 mm nut.
Load Capacity
- Pullout loads 100 - 200 % greater than plain strand (for 300 mm
lengths) depending on grout and confinement.
- Tensile capacity: Yield = 20 tonnes
Rupture = 25 tonnes.
Sensitivity
- Sensitive to nut size. Maximum recommended = 16 mm
- Sensitive to reduction of rock modulus below 10 GPa.
Strength still 2 to 3 x plain strand in these conditions
Advantages
- Reasonable ductility (adjusted through nut size)
- Small hole size
Disadvantages
- Cannot be manufactured in continuous coils at this time
- Failure mode inconsistent with large nut sizes

2.9.6 Bulbed Strand
Minimum Hole *Single: 64 mm 57 mm
Diameter (Sct.2.10): **Double: 57 mm 51 mm
*35 mm bulb used above for singles; **25 mm for doubles
Capacity Notes: (ref: Hyett et al., 1995; Bawden et al., 1995;
Garford, 1990; Stjern, 1995)
Initial Stiffness
- Approximately double the stiffness of plain strand
- 50 % pullout load after 2 to 5 mm; 100 % at 20 mm
- Limited but consistent. Wire rupture initiates at 20 to 30 mm.
Load Capacity
- Pullout loads are close to strand capacity for 300 mm
embedment lengths for W:C = 0.4 and a range of confinements.
- Wire rupture initiates at around 24 tonnes
Sensitivity
- Mildly sensitive to reduction in rock modulus below 10 GPa.
- Grout should fill bulb structure for efficient load transfer
- Increased bulb diameter 35mm= unpredictable pullout loads
Advantages
- Inexpensive and easy to install
- Can be shipped in continuous reel
- Can be easily customized to fit plates or debonded sections
- Bulb spacing and diameter can be specified to suit application
Disadvantages
- Poor grouting could lead to minimal improvement over plain
strand (Recommend W:C = 0.4).
Figure 2.9.5: Sample pullout performance of nutcaged strand
(nutcase after Bawden and Hyett, 1995; Hyett et al., 1993)
Figure 2.9.6: Sample pullout performance of bulbed strand
(Garford Pty. 1990; Bawden and Hyett, 1995; Hyett et al., 1995)
Nutcaged Strand
Bulbed Strand

Figure 2.9.7: Combinations
Figure 2.9.8: Series Combinations
a) Fault shear
b) Seismic loading
2.9.7 Combination Strand
The strand configurations on the previous
pages can be combined in parallel (Stjern, 1995)
or in series as shown at right.
Parallel combinations are double strand
cablebolt elements intended to combine the
beneficial characteristics of two different strand
types such as plain and bulbed strand. A birdcage
and plain strand combination would, for example,
also facilitate plating of a stiff modified element.
Unfortunately, in many cases, the stiffness
characteristics of the two strands may not be
compatible and the stiffer strand will rupture
before the other has a chance to carry significant
load. The capacity of the system may not,
therefore, be comparable to double strand and in
fact may be closer to that of a single strand. This
configuration is not recommended for standard
use in most mining applications.
Series combinations can be fabricated
to give different bond characteristics
along the cable. A single strand with a
debonded length (Section 2.6.4) and
with birdcaged, bulbed or buttoned end
segments would serve to provide a soft
and dynamically resilient connection
between two strong and stiff anchorages
for use in fractured ground subject to
seismic disturbance. The modified end
lengths would provide reliable bond
strength to maintain integrity of the near
face rockmass and to ensure adequate
anchorage while the debonded or plain
strand segment would accommodate
dynamic displacement or excessive fault
shear as shown in Figure 2.9.8.
2.9.8 Strand Selection
The logic for selection of strand type can be summarized as shown in Figure
2.9.9 which describes various operational and engineering considerations.
Figure 2.9.9: Cablebolt strand selection logic (modified after Windsor, 1992)
Strand Selection

Figure 2.9.10: Fibreglass cablebolts
a) Strand construction
b) Type A c) Type B
Figure 2.9.11: Bond strength vs surface type
2.9.9 Strand Alternatives: Fibreglass Cablebolts
In an effort to develop a cuttable
lightweight cablebolt which would
provide an alternative to the
conventional steel strand and its
modified derivatives, several
researchers in Canada (Pakalnis et
al., 1994; Mah et al., 1991; Mah,
1994, Peterson et al., 1992), along
with partners in the mining industry
investigated the use of fibre
composite strands (Figure 2.9.10.a).
After testing a four strand product
made from expensive European
strand (Type A in Figure 2.9.10.b)
and obtaining promising results, they
developed a more economical
product called the DAPPAM bolt
(Type B in Figure 2.9.10.c).
This bolt is composed of 10
fibreglass strands on a circular spacer
which has a grouting tube pre-
installed down the centre of the
assembly. The numerous individual
fibres in each strand are encased in a
matrix. The composite strand itself is
then coated with a surface material
which can be smooth, rough, laced
with fibre or permeated with grit.
The surface coating can be specified
when ordering and is critical to the
strength of the interface with the
chemical or cement based grout. The
effect of various surface coatings on
the pullout bond strength in Portland
cement grout (W:C=0.35) is shown
in Figure 2.9.11.
Corrosion of fibreglass
composites can be a problem in
certain environments (Reinhart and
Clements, 1988). The chemical
makeup of the matrix and surface
Figure 2.9.12: Tensile capacity comparison
Figure 2.9.13: Shear capacity comparison
(Plain strand; Windsor and
Thompson, 1993)
Figure 2.9.14: Pullout performance
coating must be selected to provide
adequate protection against alkaline
Portland cement or against the acidic
conditions found in many mines.
The Type B bolt (Pakalnis et al.,
1994) weighs approximately one
quarter as much as a single steel
strand cablebolt of the same length.
This can result in a significant
increase in productivity. Installation
trials by Pakalnis et al. have shown
comparable overall installation costs
(including drilling) of approximately
CDN $20.00/m. The bolt's main
advantage is it's cuttability. This can
lead to higher productivity in cut and
fill operations and where automated
excavators are employed. In addition,
the bolt tends to disintegrate when
blasted, removing the hazards of
steel cables in the muck stream. They
are not recommended, for the same
reasons, as support under dynamic
loading.
The Type B bolt has a design
tensile capacity of 290 kN (Figure
2.9.12) although tests have shown
breaking loads of up to 400 kN. The
shear capacity is not, however,
comparable to steel strand (Figure
2.9.13). The pullout performance of
a single Type B bolt is comparable to
a double (twin) plain steel strand
(Figure 2.9.14).
These bolts are currently available
in North America and are being used
in numerous operations. They can be
installed using similar equipment and
techniques to steel strand cablebolts.
The composite should be chemically
tested to ensure compatibility with
the milling and refining process.

2.10 Installation Configuration
The configuration of the installation includes the grout water:cement ratio,
cablebolt installation method and borehole diameter.
2.10.1 Grout Mix Design Selection
There is an optimum grout mixture for each cablebolt type which represents a
compromise between high grout strength (lower grout W:C) and good grout
flowability (higher grout W:C). Specification of the proportions of cement and
water in the grout mixture depends upon the type of cablebolt, whether plain or
modified, that has been specified in the design. A summary of the factors that
should be considered is given in Figure 2.5.9.
Plain strand cablebolts should be installed with maximum strength grout to
maximize the bond strength. Since grout strength increases with decreasing
water:cement ratio mixtures, the lowest possible water:cement ratio grout that can
be readily mixed and pumped, and which will have sufficient water for complete
hydration should be used. The grout mixture specified in design for plain strand
cablebolts should therefore be between 0.3 and 0.35 W:C. If surface fixtures will
be installed on the working end of the cablebolts, the grout water:cement ratio can
be increased to 0.4.
The performance of modified geometry cablebolts relies on grout completely
filling and supporting the cages of the bolts. To ensure that the grout will be fluid
enough to flow into the cages, a grout mix design of 0.4 water:cement ratio should
be specified in design. Wetter grout (W:C 0.4) will definitely flow into the
cages, and will be easier to mix and pump, but will have reduced strength, leading
to reduced load carrying capacity of the modified geometry cablebolts.
For additional information regarding flow properties and bond strength relative
to grout water:cement ratio, the reader is referred to Sections 2.4 to 2.6.
In summary, the optimum grout water:cement ratios are:
0.35 W:C for plain strand cablebolts,
and
0.40 W:C for modified geometry cablebolts.
2.10.2 Cablebolt Installation Method Selection
Several cablebolt installation methods are used. A brief introduction to these
methods has been made in Chapter 1, and a summary of their advantages,
disadvantages, and equipment requirements is made in Tables 2.10.1 to 2.10.3.
The selection of the cablebolt installation method depends upon the flow
characteristics of the grout mixture specified in design, and the orientation of the
borehole. The installation method must result in full encapsulation of the cablebolt
wires and complete filling of the borehole with design quality grout.
For uphole cablebolt installations, the breather tube or either one of the grout
tube installation methods can be used. Grout of 0.375 W:C will flow downward
under its own weight. Therefore this type of grout will not remain in an
unplugged uphole, and must be installed using the breather tube method. On the
other hand, thicker grouts of W:C 0.375 flow only when pumped. This thicker
grout will remain suspended in an uphole and so can be installed using the grout
tube method. Very thick grouts 0.35 cannot be effectively pumped using small
breather tubes (I.D. 12mm). Attempting to do so will result in excessive
pumping pressure, which is likely to rupture the grout tube, crush the breather
tube, or blow out the collar packing. Larger breather tubes (12-20mm) may be
used with W:C = 0.35 to 0.375 as required provided that they are always filled
with grout. The grout tube methods utilize thick grouts exclusively (W:C 0.35).
Downhole cablebolt installations are grouted using one of the grout tube
methods described in Chapter 1. Grout of W:C 0.3 to 0.45 can be specified for
downhole installations, so long as it can be easily mixed, can be pumped through
the longest grout tube, and is compatible with the cablebolt type. Grouts of W:C
0.45 are never recommended due to insufficient strength.
2.10.3 Borehole Diameter Specification
The borehole diameter specified in design must be large enough to allow easy
insertion of the cablebolt and tube(s) into the borehole, but as small as possible to
minimize drilling and grouting costs. Suggested minimum borehole sizes for
different types of cablebolt elements are shown in Figures 2.10.1 and 2.10.2.
Figures 2.10.3 and 2.10.4 can also be used to assist in the selection of the borehole
size for different combinations of cablebolts, tubes and spacers. The borehole
should be large enough to provide adequate space for easy insertion of the
cablebolt element (strand, tubes and spacers) into the borehole and grout flow
around the element. If the cablebolts stick or jam in the hole during placement,
the tubes can be kinked or crushed. In this case use a larger diameter borehole. In
a very fractured rockmass, the pieces of rock surrounding the borehole may shift,
making it difficult to insert the cablebolt into the hole. In this case, the borehole
diameter should be increased, and the cablebolt should be installed and grouted as
soon as possible after the hole is drilled.

Table 2.10.1:
Breather tube installation method (for upholes only)
Grout flow Upward against gravity. Grout flows
upward in the borehole, and then
returns to the collar of the borehole
inside the breather tube.
Grout mix
design
0.4 water:cement ratio.
Grouting
materials
1 metre of 3/4 inside diameter
(I.D.) grout tube inside the hole at
the collar + 1/2 I.D. breather tube
to the end of the borehole. Collar
packing materials are required.
Associated
cablebolt
hardware
Plain strand and all types of
modified geometry cablebolts can
be grouted with this method.
Grout pump
selection
A piston pump is commonly used
with this method. A progressing
cavity pump can also be used.
Advantages The flow of the grout upward inside
the hole, against gravity, results in
the formation of a complete grout
column without voids (except in very
fractured rock; see below).
This method should be used for
grouting multiple strand or modified
geometry cablebolts, so that the
grout will flow completely around
the wires of the cablebolts.
The breather tube is of smaller
diameter than a grout tube, and so
a smaller borehole can be drilled.
Disadvantages The encapsulation of the cablebolt
can only be guaranteed if grout
returns along the breather tube to
the collar of the hole. Blocked or
crushed breather tubes or flow of
grout away into the rockmass
through fractures will prevent the
return of grout down the breather
tube, leaving a void in the column.
The pressure required to force the
grout back down a small breather
tube may lead to burst grout tubes,
crushing of the breather tube, or
blow out of the collar packing, and
hence a partially ungrouted column.
Collar packing and borehole
grouting are usually done on two
passes, increasing labour costs.
Table 2.10.2:
Grout tube installation method
Grout flow Downholes: Upward against gravity.
Upholes: Up inside the grout tube,
and then downward with gravity
inside the borehole.
Grout mix
design
Downholes: Any water:cement ratio
(0.3 to 0.45) that is compatible with
the type of cablebolt being installed.
Upholes: 0.35 water:cement ratio.
Grouting
materials
3/4 inside diameter (I.D.) grout
tube to the end of the borehole.
Associated
cablebolt
hardware
Single plain strand cablebolts work
best with this grouting method in
upholes. Usage of this method with
modified geometry or multiple
strands may result in encapsulation
problems, in which portions of the
column or the cages are not filled.
This can be investigated with pipe
pumping tests (Section 2.12).
Grout pump
selection
A continuous feed or progressing
cavity pump should be used with
this method. Ideally a piston pump
should not be used with this
method, since the grout front
progresses in a stop-start fashion,
likely preventing formation of a
uniform grout flow front. This may
introduce voids into the grout
column or result in freeze-up of
grout in the tubes preventing
complete grouting of the hole.
Advantages Grout with W:C 0.38 will not
remain in upholes, but will run out.
The incompletely grouted cablebolt
holes should then be evident during
subsequent inspection.
The higher strength of the thicker
grout pumped with this method will
increase the cablebolt capacity.
Disadvantages Full encapsulation of the cablebolt
is not guaranteed, and may only be
partial if the grout front separates
into tongues. If the grout falls in
blobs, but remains inside the
uphole, the ungrouted hole may
appear to be full of grout.

Table 2.10.3:
Grout and Retract installation method (cable inserted before or after grouting)
Grout flow In this method, the grout flows within
the grout tube to the desired position
in the borehole. If the tube retraction
is too slow, the grout will flow a
short distance within the borehole.
Grout mix
design
Downholes: Any water:cement ratio
that is compatible with the type of
cablebolt.
Upholes: 0.35 water:cement ratio.
Grouting
materials
3/4 inside diameter (I.D.) grout
tube to the borehole end. The tube
is retracted and can be reused.
Associated
cablebolt
hardware
The authors' experience with this
method is limited to plain strand
cablebolts. However, it is thought
that the pressure created in the
grout column as the cablebolt is
inserted into the hole should result
in full encapsulation of multiple or
modified geometry strands. Full
grout encapsulation can be
investigated using pipe pumping
tests (Section 2.12).
Grout pump
selection
A progressing cavity pump should
be used with this method. A piston
pump can also be used with this
method, so long as it is powerful
enough to pump grout to the end of
the longest borehole in the pattern.
Advantages Grout with W:C 0.38 will not
remain in upholes, but will run out.
The poorly grouted, empty cablebolt
holes should then be evident during
subsequent inspection.
The higher strength of the thicker
grout that can be pumped with this
method increases cable capacity.
Disadvantages Voids can be left in the borehole if
the grout tube is retracted too
quickly, or if the grout is too wet and
falls in blobs down inside the hole.
Some grout will be displaced from
the hole as the cablebolt is inserted.
However, if grout continues to fall or
flow from the hole after the
cablebolt has been placed, there
are likely to be voids in the column.
Figure 2.10.1: Minimum recommended borehole sizes for single strand cablebolts. The
minimum recommended tube sizes are : Breather tube = 10 mm I.D.;
Grout tube = 17 mm I.D. Increase the borehole size if undue resistance
is encountered when placing the cablebolt element.

Figure 2.10.2: Minimum recommended borehole sizes for twin strand cablebolts. The
cablebolt wires shown as open circles indicate the position of the next
cage in the offset strands. Tube sizes shown are the minimum
recommended: Breather tube=10 mm I.D.; Grout tube=17 mm.
Figure 2.10.3: Cablebolt strand geometry. The cablebolt wires shown as open circles
indicate the position of the strand in the next, offset cage along the
modified geometry. Check that the dimensions shown are correct
for the specific cablebolts on the site.

Figure 2.10.4: Borehole and tube geometry. Note that in some applications, where
grout flow is unduly restricted by normal sized breather tubes, grout
tubing is used for the breather tube. Add any hole or tube sizes in
use on the site that are not shown here.
2.11 Selection of Installation Equipment
The equipment used in the installation of cablebolts has the following
requirements:
+ The drilling equipment must be able to drill boreholes of the maximum length
and diameter specified in the design. The cablebolt element (cablebolt
strand(s), spacers and tubes) must fit easily into the borehole.
+ The grout mixer must deliver well mixed batches of grout of the specified
water:cement ratio in a reasonable amount of time. The possible access
constraints of the site must be considered when selecting the pump. For
example a large, heavy mixer should not be chosen for a work place with
limited or difficult access, such as a drift only accessible by a small raise. On
the other hand, in cases where all working areas are accessible via drifts or
ramps, large twin hopper mixers could be used, as long as equipment to
transport the mixer is readily available.
+ The grout pump must be powerful enough to completely fill the longest
borehole with grout of the design water:cement ratio. As with the mixer, the
equipment must be portable enough to be easily moved into the work place
with the most difficult access at the mine site.
2.11.1 Drilling Equipment
The complete description of the specifications of drilling equipment for use in
cablebolt installations is outside the scope of this handbook. However, the
following few points indicate some of the requirements of the drilling equipment
for cablebolting.
+ The drilling equipment should be able to drill holes with reasonable accuracy.
A rule of thumb for accuracy is that the end of the holes (20 to 25 metres long)
be within 0.25 metres of the design position. The tolerance for borehole
deviation may be even tighter for cut and fill operations.
+ The equipment must be able to drill and clear cuttings from the maximum
length and most extreme angle of borehole that will be used in the cablebolt
pattern.
+ The drilling mechanism and bits least likely to damage the wall rock of the
boreholes should be used. Failure of the rock around the cablebolt hole will
reduce the confinement provided by the borehole and thus the capacity of the
cablebolt.

2.11.2 Grouting Equipment
There are numerous grout mixers and pumps available on the market with a
wide variety of options for portability, batch mixing speed, pumping mechanism
and power, ease of clean up and maintenance, and cost. The grouting equipment
selected for each mine site must be powerful enough to mix the specified grout
mix quickly and completely, and to pump the maximum length hole full of grout.
It should be possible to find the best equipment for the site by investigating the
equipment available from different suppliers. The search for the best grouting
equipment could include the products available from cablebolt, shotcrete, mining
and civil engineering suppliers. If the capabilities of the mixer and pump are not
well known for the grout consistency to be used in your application, then the
supplier should give a demonstration of the operation of the equipment, perhaps
by conducting a pipe pumping test that simulates the length and diameter of the
boreholes at the site. Instructions for pipe pumping tests are given in Section 2.12.
The term grout used in civil engineering literature usually refers to W:C ratios
ranging between 1 and 7. The much thicker grout mixes used with cablebolts
(W:C = 0.35 to 0.4) require increased mixing and pumping power and are more
sticky and abrasive, leading to more frequent replacement of machine
components and the requirement for more thorough clean up procedures.
Equipment Portability
The grout mixer and pump must be portable and robust enough to permit
frequent moves. If access to the working areas is limited, such as in some cut and
fill operations, or in captive drifts, then the size and weight of the equipment will
be limited to what the crew members can readily move by hand. The equipment
must also be robust enough to survive any possible rough handling during
transport between working areas. In mines with good access, a larger pump may
be mounted on a vehicle for easy transport and to protect the equipment from
damage.
Grout Mixer Selection Considerations
The mixer selected must be able to completely mix a given volume of design
consistency grout in a reasonable amount of time. The phrase shear mixer is
often used to describe the mechanism of a particular mixer. True shear mixing
creates velocity differentials within the grout mix, so that the grout is not simply
spun in the bin, but undergoes radial and axial convection, creating a more
thoroughly and completely mixed grout.
The grout mixers currently in use in cablebolting applications include drum
mixers, colloidal mixers and paddle mixers.
Figure 2.11.1: Effect of mixing method on grout strength from samples mixed in
the laboratory (after Gendron et al, 1992).
Drum Mixers
Drum mixing equipment is commonly used by the construction industry to mix
small volumes of concrete. The mixing mechanism of a drum mixer is similar to
the action of a clothes dryer. The rotation of the bin around a 45-axis, forces the
material to tumble under the force of gravity from one side of the bin to the other.
In concrete applications, the tumbling aggregate assists the mixing of the water
and cement. Aggregate is not used in cablebolt grout mixes however, making this
a very inefficient, ineffective mechanism for complete mixing of cement grout.
As can be seen in Figure 2.11.1, a drum mixer is likely to produce weaker grout
for a given water:cement ratio than will a paddle mixer.
Colloidal Mixers
In colloidal mixers, a small diameter, vortex rotor creates high speed, turbulent
flow. The shape of the mixing chamber or bin is carefully designed to promote
mixing as well. Complete wetting of even the finest particles of cement (colloidal
mixing) is achieved in a very short time.
Colloidal mixers have not been observed by the authors in cablebolt grouting
operations. These mixers are used for shotcrete in underground mines, and should
be readily useable in cablebolting applications.

Figure 2.11.2: Possible grout mixer bin lids which will cut and support the cement
bag. The screen is intended to remove any lumps from the dry powder.
Paddle Mixers
Paddle mixers are the most common type of mixer used in cablebolting
applications. The paddles are attached to a shaft which rotates around a horizontal
or vertical axis within the bin. If the paddles are rotated too slowly, or if the
paddles are poorly designed, they will pass through the grout mixture, not creating
any shear movement, and will produce a poorly mixed grout. Baffles may be
attached to the walls of the bin to promote shear mixing within the grout, however
dry cement or pockets of water may collect behind poorly designed baffles. The
grout should be mixed until the consistency does not change with further mixing.
The degree of mixing is especially critical when using a piston pump, since little
additional grout mixing takes place once the grout leaves the bin.
Some guidelines for selecting a paddle mixer for cablebolt grout mixing are:
+ A bin cover which cuts and supports the cement bags is very useful. The bags
can be cut by a serrated blade or cone which is placed on top of a layer of
screen. The screen should be small enough to remove any pre-hydrated lumps
of cement, but not so fine that excessive dust is created. The screen should be
positioned on the cover so that the dry cement is added to the region of highest
mixture flow velocity within the bin (away from the centre and side walls).
+ The mixer should be designed so that the grout flow moves in a turbulent
manner throughout the mixing bin, and does not spin around at the same
elevation. This can be achieved by well designed blades and baffles that are
placed at an angle to the direction of rotation. Mixing is aided by recirculation
of the grout through the pump and grout hose and back into the bin.
+ Rounded blades or flat blades aligned perpendicular to the direction of rotation
are unacceptable, since they simply pass through the grout without inducing
any internal shearing of the cement and water. Flat blades placed at an angle
to the direction of rotation should be used.
+ Blades (usually replaceable rubber strips) which scrape around the walls and
floor of the bin can be used to remove any pockets of dry cement.
+ Dead spaces within the bin, where cement powder or water can collect without
becoming part of the grout mix, should be avoided in the mixer design. Dead
spaces can be created behind baffles which are located at 90- to the direction
of the grout flow. Therefore baffles should only be used if they can be
designed to increase the shearing of the mix without creating dead spaces.
Where they are used, the baffles should be angled to the direction of the grout
flow. Dead spaces can also be created around the shaft of the mixer, where the
grout velocity is lowest. Therefore the paddles should be spaced out near the
walls of the bin and close to the shaft as well. Dry cement should not be added
along the bin walls or at the shaft.
+ The shape of the mixing bin is also important. Any sharp angle changes in the
surface of the bin, such as at the edge of any embayments, may collect unmixed
water or dry cement. Therefore the interior surface of the bin should be as
streamlined as possible to promote optimum grout flow and mixing.
+ The bin should be provided with a convenient and easy mechanism for tipping
or pouring the grout mixture into the pump hopper.
+ The bin should be easy to clean, with easy access to the paddle blades and
baffles. The interior surface of some mixer bins are plastic coated to reduce the
chance of grout sticking to the walls or base.
+ The size of the bin should provide reasonable batch volumes. See the
discussion on the following page regarding the selection of batch size.
+ The correct water:cement ratio of the mixed grout is essential for optimal
performance of the cablebolt and for use with the particular installation method.
Some mixers provide a water metering device, which can help to control the
grout water:cement ratio. Water is always placed in the mixing bin first, and
then the dry grout is added slowly.

Figure 2.11.3: Grout volume required per unit length of cablebolt borehole.
Grout Mixing
Grout must be completely mixed in batches to ensure that all of the grout
pumped into the holes is of uniform quality and strength. Mixing is complete
when there is no further change in the consistency of the grout and the surface of
the mixture appears smooth. Some mine sites advocate continuous mixing to
speed up the grouting process, however tests conducted on grout sampled from
continuous mixes at mine sites have demonstrated a wide variability in strength.
The best configuration for efficient cablebolt grouting is a double batch system
with two mixing bins that keep the pump continuously supplied with grout.
The optimum size of a grout batch can be determined through trials with the
cablebolting crew. The batch must be:
+ mixed from a full number of cement bags (usually bags are 25 or 40 kg.)
+ small enough to be easily mixed, and used up before grout set starts.
Generally, each batch should be used up within 15 minutes (0.35 W:C) to 30
minutes (0.4 W:C) from the end of mixing.
+ able to fill a complete number of boreholes. Figure 2.11.3 can be used to
estimate the volume of grout required to fill a borehole. Partial hole grouting
from a single batch should be avoided, because if the next batch is delayed, the
grout will set up in the tubes and the cablebolt hole will be lost.
+ large enough that the cablebolt crew is kept well supplied with grout. Long
mixing times will hold up the grouting operation, resulting in reduced
productivity of the cablebolting crew, and more frequent cleaning of the pump.
Grout Pump Selection Considerations
The grout pump must be able to pump the thickest grout into the longest
cablebolt borehole that will be used on the site. The grout pump manufacturers
usually provide pumping specifications for their equipment that indicate the
maximum distance grout of a specific water:cement ratio can be pumped through
a grout hose of a given diameter. The pumping specification must exceed the
most difficult pumping conditions expected on the site. If several rings of
boreholes will be grouted from a central position, the extra length of grout tube
from the pump to the furthest hole collar should be added to the required grouting
distance.
If there is any question about the ability of the pump to do the job, request that
the supplier demonstrate the performance of the pump in a pipe pumping test that
simulates the worst conditions expected at the site (longest hole, smallest diameter
tube, longest length of tube outside the hole and thickest grout).
A recirculation valve which diverts the grout flow through the grout hose and
back into the hopper is very useful. In this way continuous grout flow within the
moving parts of the pump is possible when the crew are between cablebolt holes,
or if the installation is stalled for any reason. Continued agitation or mixing of the
grout within the pump hopper will help prevent segregation or settling of the
cement particles out of the mix.
Increasing the volume of water used in the grout mixture will enable a longer
borehole to be grouted with a given pump. This is due to the reduced frictional
resistance to flow between the borehole or tube walls and the wetter grout. On the
other hand an increase in the water:cement ratio will reduce the bond strength of
the cablebolt and may prevent the use of the grout tube installation method in
upholes.
Some suppliers suggest that flushing the pump with clear water is adequate for
cleaning. However thicker grouts (W:C 0.4) may stick to the metal surfaces
inside the pump even after flushing. If the pump is not thoroughly cleaned, lumps
of hardened cement may block the grout pump tubes or installation tubes, and may
even damage the pump itself. Therefore it should be possible to take the pump
apart quickly and easily for cleaning after the grouting is complete.
Other factors to consider when selecting a grout pump are the outside
dimensions, the weight, the total operating weight, the power system, the air
requirements and the portability of the equipment.
Grout pumps commonly used in cablebolting can be classed into two groups on
the basis of their pumping mechanism: piston pumps and progressing cavity
pumps.

Figure 2.11.4: Configuration and pumping mechanism of a single-acting piston pump
Piston Pumps
The pumping action of a piston pump is shown in Figure 2.11.4. In a single-
acting piston pump, grout is pushed into the grout hose on the up-stroke alone.
Double-acting pumps push grout on both strokes. The capabilities of piston
pumps vary widely, depending upon the design of the piston, valves, and the pump
chamber, the consistency of the grout, and the power of the motor. Mine site
observations indicate that a piston pump (Spedel 6000) can fill a 12 metre long
hole with 0.4 W:C grout, using the breather tube method (Nickson, 1992). Piston
pumps are generally very portable, easy to clean and easy to maintain.
The grout batch should be agitated or mixed throughout the time that the grout
is being pumped to prevent the solid cement particles from settling out of the mix.
Some piston pumps are placed directly into the grout mixing bin, while others are
supplied with a separate bin from which the grout is pumped. In both cases, a
paddle mixer is usually used to keep the cement in suspension through out the
grouting process, and the pump is placed into a side chamber of the bin.
The pulsing action of a piston pump may prevent the formation of a consistent,
uniform flow front. If this is the case, the piston pump should not be used for the
top down, grout tube installation method for upholes.
Figure 2.11.5: Progressing cavity pumping mechanism
Progressing Cavity Pumps
The rotor and stator are the two key elements of a progressing cavity or
eccentric screw pump. The rotor is a single pitch steel spiral with circular cross
section and a degree of eccentricity. The stator is a double pitch internal spiral
with pitches at 180-. As the rotor turns, it rotates concentrically around its own
axis, and moves eccentrically as well, creating both a complete seal along the
length of the stator at all times and cavities which progress continuously (in a non-
pulsating manner) along the length of the pump.
In many of the progressing cavity
pumps currently available, the grout
is fed into the rotor/stator assembly
by a horizontal auger located at the
bottom of the grout hopper.
The pumping pressure developed in a progressing cavity pump relies upon an
absolute seal between the suction and pressure side of the pump in any position of
the eccentric screw. If the rotor is damaged, or the rubber stator begins to wear
out and enlarge, pumping pressure will be lost. At that time, replace the stator.
In field observations of progressing cavity pumps an 18 metre long uphole was
pumped full of 0.35 W:C grout using the grout tube method (Nickson, 1992).

2.11.3 Breather and Grout Tubes
The selection of appropriate breather and grout tubes is important. The cost of
the tubing is a very small portion of the total cost of a cablebolt installation, and
so the tubes are usually overlooked in the design of and material specification for
a cablebolt system.
The diameter and pressure rating of the tubes must be great enough to transmit
the grout easily. Cluett (1991) observed crushing of 6 mm I.D. breather tubes in
long holes where excessive pumping pressure was applied to try to force grout
back down an undersized breather tube. Rupture of grout tubes during grout tube
installations in long holes have been reported at a number of mine sites as well.
The selection of the tubing must take into consideration the sources of
resistance to grout flow in the tube(s):
+ The grout consistency. As the grout W:C is reduced, the frictional resistance
to grout flow increases, requiring higher pressure and larger diameter tubing.
+ The diameter and length of the tube(s). Increasing resistance to grout flow is
created by decreasing tube diameters or increasing tube lengths.
+ The number and severity of diameter changes. Frictional losses in the tube(s)
can be severe due to diameter changes along the grout flow path as is discussed
in Section 2.5.4.
Breather tubes of 11 mm inside diameter (I.D.) and 1.5 to 2 mm wall thickness
should be adequate for most cablebolt holes pumped with grout of W:C 0.4. In
Australian mines breather tubes are often 17mm I.D. Failure of the grout to return
back down the breather tube could indicate that the breather tube is too small or
that the grout is flowing away into fractures in the rockmass. In the first case, the
increase in pressure required to force grout into too small a tube will cause the
pump to stall, the collar packing to blow out, or the breather tube to crush or
collapse. In this case, increase the diameter of the breather tube. If there is any
evidence of crushing of the breather tube, increase the required tube pressure
rating. Loss of grout into a fractured rockmass is indicated when an excessive
volume of grout has been pumped into a hole, and there has not been any return
along the breather tube. In this case, use the procedures for grouting in fractured
rock given in Chapter 3.
Grout tubes are usually 17 to 25 mm I.D. and 2 to 3 mm wall thickness. The
minimum recommended pressure rating for grout tubing is 100 psi. The grout
tube may burst in situations where pumping pressure increases as thick grouts are
pumped along long lengths of grout tubing or when grout is difficult to pump back
down a breather tube (tube too small or grout too thick). In this case, higher
pressure rated tubing should be purchased.
Figure 2.11.6: Effect of fully grouted and
empty breather tubes on cablebolt capacity
(after Goris, 1990)
The occurrence of problems may
prevent complete grouting of the
borehole. The ungrouted sections of
the cablebolt strand will have no load
carrying capacity.
An open, ungrouted 11 mm I.D.
breather tube within the grout
column (0.45 W:C) has been found
to reduce the capacity of a plain
strand cablebolt by 30% (Goris,
1990: laboratory pull tests conducted
on 25 cm long samples confined in
steel pipe). On the other hand, Goris
found that fully grouted breather
tubes do not reduce the capacity of
cablebolts.
2.11.4 Installation Accessories
Installation accessories include items which are used with the cablebolt strand
itself, to help insert the cablebolt into the hole, to support the cablebolt prior to
grouting and to block the collar of upholes.
Cablebolt Dispensers
Coils containing a complete uncut length of cablebolt strand are often the
cheapest and most convenient way to purchase the strand. If it is properly handled
and stored, a coil will provide a large supply of clean strand. Coils of cablebolt
strand may be shipped in individual packs which are bound with steel straps, or
within refillable dispensing cassettes. Mine sites which use individual coil packs
usually manufacture or purchase their own refillable racks or dispensing cassettes.
Dispensing cassettes or racks can be horizontal or vertical, and stationary or
rotating. The dispenser should be easy to load. Whatever the configuration of the
dispenser, the prime requirement is that the cablebolt does not become tangled and
unwound. Guides through which the cablebolt is pulled may help to keep the coil
together and untangled. The strand is usually pulled from the centre of the coil in
stationary dispensers. The strand will acquire a twist for each loop that is removed
from the coil. The twist should be in the direction of the cablebolt lay, so that the
cablebolt strand will tighten up instead of unwinding. The strand is usually pulled
from the outside of the reel of a rotating dispenser. A brake to stop the rotation
of the dispenser should be provided.

Figure 2.11.7: Possible cablebolt dispensers
Figure 2.11.8: Cablebolt hangers
Cablebolt Hangers
The cablebolt must be well secured in uphole installations, between the time of
placement of the cablebolt and the time of grout set. Some examples of hangers
observed in use in underground mines by the authors are shown in Figure 2.11.8.
The most important consideration when selecting the hanger or collar wedge for
use at a site is that the cablebolt must remain completely secure in the uphole until
the grout has set. A cablebolt falling from an uphole can cause serious injury.
The choice of a particular hanger will depend upon the ease of use and the cost of
the hanger, the borehole diameter, the method of inserting the cablebolt in the
hole, the grouting method, and prevention of damage to the cablebolt strand and
tubes. Longer cablebolts will require stronger hangers. For example extra steel
strips should be added to spring steel hangers for long cablebolts. In situ pull tests
on cablebolts with hangers can be used to ensure adequate hanger strength.
There are a number of different configurations of hangers and wedges in use
at mine sites. Some hangers are formed by bending one wire of the strand, while
others are separate items which are attached to the cablebolt. Hangers can be used
at either the collar or toe of the hole, and wedges are used at the collar. Toe
hangers will help support the cablebolt as it is being inserted into the borehole.
Hangers at the collar end of the cablebolt will be easier to insert into the hole, but
will not help support the cablebolt during placement.
Some hangers centralize the cablebolt in the borehole, while others push the
cablebolt to one side of the hole (offset hanger). In uphole grout tube installations,
it is advantageous to use a centralizing hanger to start the cablebolt in the middle
of the hole to promote optimum grout encapsulation around the steel strand.
When offset hangers such as the bent wire hanger are used, attach tubes that
extend to the toe of the hole beneath the overhanging, protective hanger wire.
The diameter of the borehole can be critical to the strength and support
capabilities of the hanger. If the borehole diameter is too large, the hanger will
not lodge into the rock as well and may not be able to support the cablebolt. If the
borehole diameter is too small, it will become difficult to insert the cablebolt into
the borehole.
When using wooden wedges at the borehole collar, it is very important that the
grout and/or breather tube are not pinched or constricted in any way as this will
impede the grout flow.
Hangers made by cutting one wire of the cablebolt strand have deliberately not
been included in Figure 2.11.8, because these hangers reduce the capacity of the
steel strand (by at least 1/7 or 15%) and lead to eccentric loading of the cablebolt.
In addition, there are very few tools available which will cut just one wire of the
strand, so the capacity reduction of the cablebolt at the position of the hanger can
be much greater than 15%.

Cablebolt Spacers
Cablebolt spacers are often used in Australia when more than one cablebolt is
installed within a single borehole. The spacers keep the cablebolts separated and
away from the borehole wall, aiding complete grout encapsulation of the strands.
In some circumstances, where a single cablebolt should be centralized within the
borehole (such as in corrosive environments in which the grout provides protection
against deterioration of the steel), spacers may also be used.
Spacers are usually formed of plastic and so are relatively inexpensive. The
spacer must snap firmly onto the cablebolt strand(s) so that it will remain in place
as the cablebolt element is inserted into the borehole. The edges of the spacer
should be rounded to prevent snagging on the borehole wall. The borehole
diameter should be 5 mm greater than the largest cross-sectional dimension of
the spacer. The cross-sectional shape of one spacer from Australia is shown in
Figure 2.10.2.
Borehole Collar Sealing
Collar sealing is required for uphole grouting installations. The collar can be
sealed on a first pass or at the time of grout pumping. Collar sealing materials that
are commonly used on a first pass include burlap, cotton waste, expansive foam
and grout plugs. Collar sealing can also be done at the time of grouting with a
rubber cone or victaulic pipes. Drawings and installation procedures for each of
these collar sealing methods are given in Chapter 3.
Grout Tube Connectors
Connectors are required to join the grout pump hose and the installation grout
tube during grout pumping. The connector must be easy to fasten and strong
enough to withstand the pressure built up in the grout hoses. The most common
type of connector consists of a plastic screw cap with a hole in the centre which
fits over the grout tube. This cap screws onto a threaded metal or plastic end
which is attached to the grout pump hose. Some mine sites have also used vice
grips with a modified circular grip to hold the two tubes together.
Cablebolt Cutters
The cablebolt strand supplied in a coil is cut into lengths prior to installation in
the borehole. Cablebolt cutting is also required at the face in cut and fill stopes
where lengths of the strand are exposed by successive mining lifts. Several
methods can be used.
An air powered grinder will cut through a cablebolt strand fairly quickly if the
grinder blade is replaced as soon as it gets dull. The grinder can be a hand tool or
where possible may be a larger tool which is permanently mounted on a truck or
on a working platform in a central stores area. The production of sparks and steel
shards when cutting the cablebolt necessitate the use of a face shield and leather
gloves by the operator.
An oxyacetylene torch will very quickly burn through the cablebolt strand, but
requires a lot of bulky materials including the cutting torch, regulators, hoses and
the oxygen and acetylene bottles. The use of pressurized gases and flame will be
a safety concern in some mines as well. In addition, the operator must be careful
to avoid touching the hot end of the cablebolt for a period of time after cutting.
Hydraulic cutters can be very efficient if they are large enough to cut through
the cablebolt strand in one action. The large size and weight of a powerful
hydraulic cutter make it practically useable only when it can be mounted on
mobile equipment. In this case, the operator can be far enough away from the
cablebolt to be removed from any shards or whipping of the cablebolt.
Explosives have been used to cut cablebolts exposed in cut and fill mining.
Explosives suitable for cutting cablebolts are available from some suppliers. It
must be possible to attach the explosive to the cablebolt safely and for the
explosive to cut the strand easily. In addition, the explosive must be readily
available and safely useable in the underground working environment.
Cablebolt Pushers
Cablebolt pushers consist of two rotating rollers between which the cablebolt
passes. The surface of the rollers should be formed of strong material that will
resist tearing and undue wear. The rollers must grip the cablebolt tightly enough
and have enough power to push the longest cablebolt into an uphole. Cablebolts
slipping within under-designed pushers have been reported at some mine sites. On
some pushers, the rollers are individually powered by air motors so that if the
cablebolt jams in the hole, the motors will stall and there will be no damage to the
unit. For pushers used to place modified geometry cablebolts, the rollers may be
mounted on spring loaded arms to accommodate the changes in the diameter of the
cablebolt. A pusher mounted on an articulated arm can be easily positioned
beneath the borehole collar.
Cablebolt Trucks
The cablebolting operation can be streamlined when a well equipped truck is
supplied to the crew. The utility of a cablebolting truck depends on good access
to the working areas. Depending upon the specific requirements of each site, the
truck can be outfitted with a scissor lift, cablebolt cutter, cablebolt pusher, strand
storage platform or coil dispenser, grout mixer and pump, grout storage platform,
water supply and lights.

2.12 Pipe Pumping Test Procedures
Pipe pumping tests can be used to investigate the completeness of grout
pumping into the cablebolt hole or the capabilities of the equipment, and for
training the crew in cablebolt installation procedures. The advantage of pipe
pumping tests is that the grout column can be cut apart after the grout has set to
investigate the quality and completeness of the grout column.
The materials required for pipe pumping tests are:
1) Pipes. The number, length and diameter of the pipes required will depend
upon the objectives of the test, but are usually of the same length and
diameter as the cablebolt holes. The pipe material chosen should be robust
enough to survive the conditions of the test, but also be within the test
budget. Steel pipes are strong, but opaque and hard to cut; PVC pipes are
strong but opaque; and clear acrylic pipes are transparent so that the grout
flow front can be observed during the test, but brittle and expensive.
2) Pipe joiners for longer test lengths. The joiners should fit over the outside
of the pipes and must not impede the flow of the grout.
3) End caps to seal the toe end of the pipes.
4) Grout mixing and pumping equipment.
5) Cement and any additives.
6) Cablebolt strand(s).
7) Breather and /or grout tubes.
8) Hangers, collar wedges, spacers and collar packing materials.
9) When required, cylinders for collecting grout samples for strength or
sedimentation tests.
The test procedure should follow these steps:
+ Suspend or support the pipe in the orientation of borehole.
+ Place the cablebolt following normal procedures as laid out in Chapter 3.
+ Mix and pump the grout, ensuring that the grout mix proportions are correct.
Note the time at which grouting starts.
+ During transparent pipe tests, monitor the flow of the grout within the pipe and
note any areas where flow is unduly constricted, or where some of the column
is ungrouted. A video and/or photographic record of the tests is very useful for
training purposes and for analyzing the test results.
+ Make a note of the time that grout first appears at the end of the breather tube
or at the collar. If this grout has a watery appearance, keep pumping until grout
of the mix consistency begins to flow from the tube or hole collar. Record the
time that the pumping continued after the first appearance of grout.
+ Kink over and tie off the tube(s).
+ Make detailed notes about each test, including information about the W:C,
mixing time, pumping rate, progress of the grout flow front, and any problems.
Figure 2.12.1: Grouting problems found in pipe pumping tests conducted by the authors
and at a number of mine sites
To examine the completeness and quality of the grout column, cut sections
through the pipe and cablebolt, after the grout has set ( 72 hours). If a modified
geometry cablebolt was tested, some of the slices should be made through the
flared sections and some through the sections of regular strand.
Some problems which have been observed in pipe pumping tests conducted at
mine sites and by the authors are shown in Figure 2.12.1. Photos of sections cut
through pipe pumping test samples are given by Goris et al. (1994). Empty voids
in the grout column have the potential to reduce the load transfer capability and
hence the capacity of the cablebolt in their vicinity to zero. The problems shown
here will not always occur at each site, but should be considered if the cablebolts
at a site do not perform as expected. Other quality control problems, and the
potential influence of stress change on the cablebolt bond strength should also be
considered in any investigation of poor cablebolt performance.
If any empty voids are found during the tests, or if incomplete grouting of the
cablebolt holes are suspected, the problems could be due to:
+ poor centralization of the cablebolt in the hole,
+ incomplete grout mixing,
+ inappropriate grout water:cement ratio,
+ inappropriate grouting method, and
+ inadequate breather or grout tube size.

Figure 2.13.1: Demand assessment for support design in underground mines
2.13 Demand
While capacity determination answers the question What can be done ?,
demand assessment is the process of determining What should be done ?.
Design relates these two questions to arrive at a solution which satisfies both,
within the constraints of economics, operational limitations and practical ability.
2.13.1 Excavation Response
Figure 2.13.1 illustrates the basic flow of investigation involved in rigorous
demand assessment. The first goal is the determination of rockmass properties and
excavation influences. Then, either through empirical (experience based) methods
or though mechanistic analysis of rockmass behaviour, stability and support
demand can be evaluated. The following sections form a practical summary of this
methodology. Readers interested in more detailed rock mechanics are directed to
Bouchard (1991), Brady and Brown (1985; 1993), Goodman (1976; 1980), Hoek
et al. (1995), Hoek and Brown (1980), Hudson (1989), Jaeger and Cook (1979).
Two relationships are central in determining general excavation response
characteristics (Figure 2.13.2):
+ The ratio of far field or induced stresses to rock strength
+ The ratio of block size to excavation dimension
Stress (crudely defined as loading over an area) results from rock loads acting
at depth and from tectonic adjustments within the earth. The stress field (the
variation of stress with orientation over three dimensions) is disturbed by the
creation of underground openings. As a result induced stress changes occur which
can relax the rock mass (destress) or which can increase the stresses tangent to the
boundary to a point close to the strength of the rock. In the latter case, fracture,
damage and ultimately disintegration can occur. Stresses are either compressive
(tending to push in) or tensile (tending to pull apart) and have their respective
counterparts of compressive and tensile strength. Shear (distortional) stress results
from oblique combinations of the above normal stresses in three dimensions.
Strength is the ability of the rock to withstand elevated levels of stress without
sustaining damage. Yield strength marks the onset of such damage while ultimate
strength indicates the maximum limit of stress which can be endured before
complete rupture. Failure is an arbitrary term which must be qualified with respect
to these limits; yielded rock may carry significant load around an excavation.
Block size relates to the average dimension of competent rock blocks created
by the intersection of natural breaks in the rock called discontinuities. Joints are
natural extension features, while shears are discontinuities which reflect previous
or ongoing movement (relative slip) between two blocks of rock. All rockmasses
possess some discontinuities. The influence of these features on stability of an
opening is normally controlled by the relative inter-joint spacing or block size with
respect to the dimensions of the excavation. Clearly, rockmasses with closely
spaced discontinuities (Heavily jointed, fractured or broken rockmasses) are more
likely to have stability problems and to require the addition of artificial support,
than are relatively intact or massive rockmasses. A third scenario develops when
the rockmass has a moderate concentration of highly persistent (long)
discontinuities which mutually intersect in the vicinity of an opening. These joints
or shear structures can form large intact blocks, slabs or wedges which can be
released into the excavation and which may require support.

Figure 2.13.2: Stress, structural integrity and failure modes (after Hoek et al., 1995)
Rockmass Behaviour
Figure 2.13.3: Stress
Figure 2.13.4:Stress on a
a) Plane b) Volume
Figure 2.13.5: a) Elemental cube showing tensor convention (+ve directions shown).
b) Sections through elemental cube c) Principal stresses with typical notation
2.13.2 Stress - A Brief Introduction
Stress in its simplest form can be calculated in a one-
dimensional example (Figure 2.13.3), as load divided by
the area over which the load acts. Stress acting on a plane
can have two components. One is a normal component
acting perpendicular to the surface. The other is a shear
component acting parallel to
the surface (Figure 2.13.4a).
When acting on a separation
plane between two solid masses, a normal compressive
stress will tend to push the two halves together (a
negative or tensile stress will pull them apart). A shear
stress acting on the separation plane will tend to slide
the two halves past each other in opposite directions.
Normal compressive stresses acting on a solid will
compress or collapse the solid (compressive strain) as
shown in Figure 2.13.5. Shear stresses will cause an
angular distortion (shear strain) as shown.
Stresses in 3 dimensions are more difficult to visualize. Within a rockmass at
depth, stresses act in all directions upon a sample unit volume and are associated
with three dimensional deformation of the unit volume (strain). These stresses
(and the corresponding strains) vary with direction. The mathematical entity used
to describe such a state is the stress tensor which expresses the three normal
stresses and six shear stresses acting on the faces of a fictitious and infinitesimally
small cube (placed within the stress field) in three orthogonal directions at
specified orientations. While the stress state at a point is unique, the tensorial
description depends on the orientation of these reference axes. Figure 2.13.5.a)
shows a schematic representation of a stress tensor expressed with respect to the
global axes shown. The three sets of coplanar stress components are illustrated in
the sections in Figure 2.13.5.b). Note the equality of co-planar shear stresses.

Figure 2.13.6: 2-Dimensional stresses; tensor rotation and principal stresses.
For every valid stress state, there exists a unique orientation of the reference
cube (Figure 2.13.5.c) at which the shear stresses reduce to zero. The remaining
normal stresses are called the principal stresses and are usually quoted as major
principal stress . , intermediate principal stress . , and minor principal stress .1 2 3
(compression positive sign convention). The magnitude and orientation of these
principal stresses uniquely defines the state of stress at a point. The orientations
are quoted as trend (angle CW from North) and plunge (downward angle from the
horizontal).
The concept of tensorial stress and principal stress directions and magnitudes
is easier illustrated by considering only one of the sections in Figure 2.13.5.b) and
ignoring all out of plane stresses (Figure 2.13.6). The two-dimensional stress state
shown does not change as the square reference element is rotated. Only the tensor
description changes - hence both the tensor magnitudes and reference orientation
must be specified when describing a stress state in this way. For a certain
orientation (at the right of Figure 2.13.6) the shear stresses vanish. The resultant
normal stresses are the principal stresses (in 2-D denoted by . and . ). The1 3
orientation of the principal axes must be specified when quoting principal stresses.
Note that this kind of two-dimensional analysis is not normally valid in a three
dimensional world. The individual components or directions of a three-
dimensional tensor cannot be considered separately and independently in this
fashion. This is the nature of a tensor. Two-dimensional analysis is only valid in
circumstances where the excavation geometry is long in one direction (e.g. a
tunnel) and one of the principal in situ stress components is aligned with the tunnel
axis or out-of plane direction. In this case, there are no out-of-plane shear stresses
and the analysis need only consider the in-plane components. The out-of-plane
direction is confined (no induced strain) in this plane-strain case. The out-of-plane
normal stress will change but can be ignored in most elastic analyses (no rock
yield) that meet the geometric conditions.
Figure 2.13.7: Calculation of stresses on a plane a) 3-D; b) 2-D
Stresses on a plane
There are applications where it is necessary to evaluate the stresses on a plane
such as a joint or fault within a stress field, in order to assess the potential for slip
or dilation. Figure 2.13.7.a shows the tensor arithmetic necessary to do so in three
dimensions. Figure 2.13.7.b gives the procedure for two-dimensional calculations.

Figure 2.13.8: Induced stress flow around an excavation; stress flow for an isolated
a) vertical stress; b) horizontal stress; c) Excavation disturbance of 2-D stress field
In Situ and Induced Stresses - Modelling
Before modelling, it is essential to establish the existing state of stress. The
initial stresses can be due to sedimentation or volcanism as well as tectonic
movements and pressures. The stress state could have been further modified
through folding, faulting, metamorphism, or erosion. Typical vertical stresses in
MPa range from 0.025 to 0.03 times the depth in metres. In a tectonically inert,
non-eroded basin (e.g. South African gold fields) the horizontal stresses at depth
may be one-half of the vertical. Near surface, where horizontal stresses are locked
in while vertical stresses are relieved by erosion of overburden, or in areas of high
lateral pressure (e.g. Canadian Shield) the horizontal stresses can vary from 1.5 to
more than 3 times the vertical stress. All three principal stresses may be
significantly different and may vary from location to location. Regional structure
(faults and dykes) will often cause significant stress field disturbance. While
several general guidelines for stress estimation are available (Hoek and Brown,
1980; Herget, 1988, Hoek et al., 1995; Zoback, 1992), it is still prudent to obtain
a local measurement of in situ stress (Herget, 1988). This initial stress tensor will
completely determine the induced stresses obtained through excavation analysis.
Excavations disturb the in situ stress field. An analogue to induced stress flow
around an excavation is the flow of river water around a bridge post. The stresses
build up (concentrate) on the sides of the excavation parallel to direction of .1
(Figure 2.13.8.ab) just like the water around the post. Where the difference
between major and minor principal stress is high, relaxation zones or even tension
may develop on the faces which lie parallel to . Typical contours of maximum3.
induced principal stress are shown in Figure 2.13.8c. Note the convention for
stress trajectories. The small ticks on the crosses represent the direction of .3
while the long ticks represent . . The stresses at right angles to and adjacent to the1
walls of the excavation are zero in the absence of support or an internal pressure.
Figure 2.13.9: Influence of far-field stress orientation on induced stresses (example)
Figure 2.13.10: Zones of overstress and relaxation. Both may lead to rockmass failure.
Hoek and Brown (1980) present a set of charts for stress analysis of simply
shaped excavations. Analysis of induced stresses around more complex openings
requires the use of numerical models. The example in Figure 2.13.9 illustrates the
effect, on the induced major principal stresses, of different in situ (far-field or pre-
excavation) stress orientations. Figure 2.13.10 shows both major and minor
principal stresses due to a more complex stope geometry. Excavations often
possess zones of overstress and of relaxation. High stress may cause rupture while
low stress allows joints to dilate and blocks to unravel under gravity loading.

σ σ σ σ σ σ1 3 3
2
S C C Cm s UCS m s= + + =where and are material constants.,
2.13.3 Strength
Stresses tangent to and adjacent to the face of an excavation can be compared
directly to the uniaxial compressive strength of the rock or rockmass (Section
2.14.2). The laboratory values, however, usually overestimate the strength in the
field. Typically, field strengths of visibly competent (strong and brittle) rock
adjacent to an excavation have a uniaxial or unconfined compressive strength
(UCS) of 1/3 to 1/2 of the laboratory UCS (Martin et al., 1993; 1994; 1995; Brace
et al., 1966; Bieniawski, 1967). When the tangential stresses at the boundary
exceed this value, induced fracturing of the rock may be predicted. The strength
of massive unfractured rock in tension (. ) is typically 5 to 10% of the UCS.T
The compressive yield strength of the rockmass within one excavation radius
of the boundary can be estimated by calculating the principal stress difference (. -1
. ) and comparing this value to 1/3 to 1/2 UCS. Rock zones where the induced3
principal stress difference exceeds this value may sustain damage. The rock may
only fail (fall down), however, if the confining stresses are relaxed during
subsequent mining steps.
While the initial yield or damage threshold of the rock may be given by either
of the above simple relationships, the ultimate rupture of this yielded rock may be
more dependent on confinement (Figure 2.13.11.a). Hoek and Brown (1980; 1988)
developed the Hoek-Brown criteria which relates strength in terms of maximum
allowable . to confinement in terms of . through the relationship:1 3
For intact rock, s, is given as 1.0. Note that the UCS shown here refers to the
intact rock strength. Typical values for UCS and m of intact rock specimens arei
given in Table 2.14.1. For more fractured, damaged or jointed rockmasses, Hoek
et al. (1995) present values of m/m and s for varying deviations from ideal intacti
conditions. These values are summarized in Table 2.15.1. The value of s tends to
zero as the disintegration or disturbance of the rockmass becomes complete.
Elastic (non-yielding) modelling programs will typically give factors of safety
against failure (normally the ratio of strength to stress) given the appropriate
strength parameters. Areas which show low factors of safety (especially below
1.0) are areas where support may be required, both to reinforce the failing
rockmass and to hold up the failed material against gravity. For modelling
programs which allow yielding in the analysis (inelastic or plastic analysis), peak
(initial failure, A) and residual (post peak strength, B) values may be specified as
in Figure 2.13.11.b). The residual values determine the ultimate strength of the
failed material. While ductile material (A-A') carries stress after yield, very brittle
material with little strength after yield will clearly not be able to support itself. As
well, if the rock is very strong but brittle and if the stresses are high, the system
is capable of releasing a great deal of energy upon rupture resulting in rockburst
conditions (Hoek and Brown, 1980).
( )τ σ φS nc i= + +tan
σ σ σ σ
φ φ
σ σ σ σ
1 3 1 3
1 3 1 3
2 2
2 2
−
=
+
+
− +
sin cosc
deviatoric stress confining pressurewhere is the and is the .
Joints (continuous weakness planes) can dilate and separate under low stresses
or relaxation and can also slip under gravity loading or under excavation induced
shear stresses. The tensile strength of a persistent joint surface is zero. The Mohr-
Coulomb criterion relates shear strength, J , to a confinement (normal stress, F )s n
independent strength component or cohesion, c, and a confinement dependent
component defined by friction angle, N. An incremental frictional component
(dilation angle, i) can be added to account for joint roughness (Patton, 1966):
{Figure 2.13.11.c}
Normally c is conservatively assumed to be zero. N can vary between 15
degrees (clay gouge) to 35 degrees for coarse grained rocks. Dilation angle i can
be as high as 20 degrees, for rough surfaces at low confinement, to zero at high
confinement (high F ). Alternative strength envelopes for joints are shown inn
Figure 2.13.10.c). An observed non-linearity in shear strength with increasing
confinement is analogous to decreasing the dilation angle ,i, and increasing
cohesion, c, with increasing normal stress. This simulates the tendency for
shearing through rough asperities as normal confinement increases. Barton et al.
(1973; 1976; 1977; 1990) have developed a more rigorous non-linear relationship
which is summarized in Stacey and Page (1986) and in Hoek et al. (1995).
The Mohr-Coulomb criterion can also be applied to the rockmass as a whole.
The strength envelope defines the yield point of the critical fictitious shear plane
within the rock. If half circles are plotted on the F axis between F and F , then 1 3
same equation as above can be used to find the strength envelope which is tangent
to the suite of failure circles obtained from testing data (Figure 2.13.11d).
Expressed in terms of F and F the limiting Mohr-Coulomb yield envelope (Desai1 3
Siriwardane, 1984) becomes:
2.13.4 Block Size and the Influence of Scale
The stability of excavations in any rockmass decreases with increasing scale
(excavation span). This is due in part to the involvement of larger potential zones
of rupture or of larger agglomerations of blocks. The same support density in a
larger excavation may not be as effective as in a smaller excavation. In addition,
larger volumes of rock have a higher probability of including discontinuities,
fractures and other flaws. Laboratory samples, for example, are by necessity taken
from intact pieces of drill core, avoiding cracks and flaws where possible. The
strengths obtained, then, will inevitably be higher than those achieved in situ.
Similarly, boreholes are smaller than caverns, and thus will have less chance of
intersecting critical flaws, thereby exhibiting higher strengths (Figure 2.13.12).

Figure 2.13.12: Scale dependent rockmass strength and structural integrity
Figure 2.13.11: a) Intact granite example: Initial damage and peak Hoek-Brown strength
criteria (after Martin, 1995); b) Ductile vs brittle post-peak behaviour; c) Mohr-Coulomb
and Patton shear strength - joint slip; d) Mohr-Coulomb strength for rock and rockmasses
Figure 2.14.1: Basic components of a rockmass classification scheme
2.14 Rockmass Classification
One of the most potentially complex tasks assigned to a rock mechanics
engineer is the determination of representative mechanical properties of a
rockmass. While tests have been devised to quantify strength, stiffness and other
properties of laboratory rock specimens, it is a much more daunting task to
evaluate the quality and expected behaviour of a rockmass in the field.
Fortunately, numerous researchers have developed empirical methods (based on
numerous case histories) to quantify the relative integrity of a rockmass and
thereafter to estimate mechanical properties for excavation and support design.
These methods are referred to as rockmass classification systems.

2.14.1 Rockmass Classification Components
Rockmass behaviour is controlled by the following components (Fig. 2.14.1):
Intact Rock Strength
Stronger rocks are more likely to be stable in general conditions than weaker
rocks. Stiffness usually correlates directly with strength (Deere, 1968).
Field Stresses
At moderate depth the rockmass is likely to be confined and held together
(clamped). Near surface, in late stage mining areas which have become
relaxed, joints can open up, decreasing stiffness, strength and stability. At
greater depth, stresses induced by the creation of the excavation may exceed the
strength of the rock, resulting in induced fracturing and instability.
Fracture Density or Drill Core Quality
Diamond drill core from geotechnical or exploration drilling provides a
convenient means of assessing the structural integrity of the rockmass prior to
excavation. Numerous breaks in the core indicate a highly fractured or jointed
rockmass which is more likely to be unstable when excavated.
Joint Persistence
Joints which are highly persistent (long) are more likely to combine with other
structures to form large free blocks of rock, than are short joints. These blocks
may require support to ensure stability.
Joint Spacing
Closely spaced joints result in a smaller block size, increasing the potential for
internal shifting and rotation as the rockmass deforms, and reducing stability.
Joint Contour, Aperture and Surface Condition
Planar joints are able to slip more readily than wavy or undulating surfaces.
Similarly, smooth or polished surfaces have lower frictional slip resistance than
rough or stepped surfaces. Open joints or infilled joints are less stable than tight
or healed fractures.
Groundwater
Groundwater can destabilize an excavation by eroding or weakening joint
surfaces and infillings. In addition, water pressure reduces the frictional
resistance to slip along fractures and further destabilizes the rockmass.
Joint Orientation
Joints can intersect an excavation at unfavourable orientations, creating the
potential for slabbing, sliding blocks, stress induced slip or wall separation.
2.14.2 Data Collection
Certain basic input is required for rockmass classification and for other forms
of stability assessment. Only the most common and necessary are described here.
Refer to Hoek et al. (1995), Hoek and Brown (1980), Brady and Brown (1993),
Bieniawski (1989) and other rock mechanics or rock engineering texts for
additional investigations and analyses.
Intact Rock
Uniaxial Compressive Strength, UCS
UCS is defined as the maximum uniaxial (one
dimensional) compressive stress sustained by a
cylindrical sample, in a laboratory test (I.S.R.M.,
1981), before disintegration (as shown at right).
The stress is calculated as maximum load divided
by the cross-sectional area of the sample.
Elastic Stiffness - Young's Modulus, E
Young's modulus, E, the modulus of elasticity
defines the slope of a linear approximation of the
response curve (stress vs strain) at stress levels
around one-half of the uniaxial compressive
strength. In this region, the sample is assumed to
be elastic and the strains (deformation / length)
are assumed to be fully recoverable upon
unloading. Poisson's Ratio, , is a third important
parameter for numerical analysis. In the sample
test described above, relates the radial strain to
the axial strain for a given stress increment.
Table 2.14.1: Typical Intact Properties (after Stacey and Page, 1986; Hoek et al., 1995)
Rock Type UCS
MPa
m i
s=1
E
GPa
Rock Type UCS
MPa
m i
s=1
E
GPa

Andesite 240 19 60 0.2 Gneiss 220 33 60 0.2
Basalt 230 17 60 0.2 Granite 220 33 60 0.2
Diabase 240 19 90 0.2 Limestone 180 8 70 0.3
Dolerite 240 19 90 0.2 Sandstone 40-80 19 20 0.2
Dolomite 100 10 70 0.2 Shale 120 * 4-9 15 0.1
Gabbro 280 27 90 0.2 Quartzite 240 24 80 0.2
* with marked anisotropy (strength varies with loading direction across laminations)

In Situ (Far Field) Stresses
The determination of the local in situ stresses in the area of a proposed
excavation is beyond the scope of this book. Consult local databases and refer to
Hoek et al. (1995), Herget (1988), Brady and Brown (1993) and other texts for
general stress-depth relationships. It is preferable however, to measure the local
stress field using special instruments and procedures described in the above texts.
Structural Data
Again, Hoek et al. (1995) and Hoek and Brown (1980) and Hoek and Bray
(1981) and others describe field mapping techniques used to evaluate the structural
integrity of the rockmass. It is first necessary to identify the presence of major
through-going structures such as shears, faults or major weakness zones in the
vicinity of the proposed excavation. These discrete structures must be assessed
separately as they will dominate local behaviour.
Next, the ubiquitous (present everywhere) structure must be assessed. Systems
of extension joints and minor shear structures will have formed under historical
stress fields, which were relatively consistent over a local region. As a result, there
are usually several distinct groups of similarly oriented structures within a
rockmass. These are called joint sets. Ungrouped joints are referred to as random.
A compass (Figure 2.14.2a) is typically used in the field to record the
orientation of joints in the wall, floor or back. In addition (Figure 2.14.2b) it is
necessary to record qualitative information about the joint surfaces for rockmass
classification and later analysis. A minimum of 100 local measurements are
normally required to define the structure in a zone of rock. More measurements
improve the data reduction accuracy. It is necessary, however, to restrict the data
to distinct, local groups in areas of changing rockmass quality and nature.
The stereonet (Figure 2.14.2c) is used to visually and statistically resolve the
data into clusters or sets. Computer software such as DIPS (Diederichs and Hoek,
1989; Hoek et al., 1995) can be used for this purpose. Representative (mean)
orientations for each cluster are used in analysis.
Discontinuity Strength
Refer to Barton and Choubey (1977) or Hoek et al. (1995) for detailed joint
strength determination and application. For simplified Mohr-Coulomb analysis
(c=0, N, and i in Section 2.13.3) near excavation boundaries, friction, N, varies
from 20 degrees for schistose joint walls, to 30 degrees for competent granular or
crystalline rocks. Add 5 degrees if the joint is completely dry. The dilation angle,
i, for low confinements, varies from 2 degrees for smooth joints, to 6-10 degrees
for rough joints and to more than 14 degrees for very rough or stepped surfaces.
Figure 2.14.2: Joint mapping and orientation analysis (after Diederichs, 1990)
Structural Data

Figure 2.14.3: Conventional method for evaluating RQD from drill core
2.14.3 Rock Quality Designation, RQD
Deere et al (1967) developed the Rock Quality Designation, RQD, in response
to the need for a quick and objective technique for estimating rockmass quality
from diamond drill core logs during the initial exploratory phase of construction.
The method is simple and efficient to implement in mining environments and can
be assigned to the drillers themselves or to the geologists analyzing core for grade
assessment. Information gained at this early stage of exploration is extremely
valuable for the geomechanics engineer involved in the mine planning process.
RQD is calculated as the ratio of the sum of the lengths of all pieces of core
greater than 10 cm to the total length of the core run. This total length of core
must include all lost core sections. Breaks created by the driller during removal
from the core barrel are to be ignored. Core discing due to high stress should not
be considered in the calculation of RQD but should be noted in the drill log (give
an estimate of discing frequency). Discing does not contribute to RQD but does
indicate potential risk of brittle overstress problems during excavation.
A great deal of work has been done to correlate RQD with joint frequency,
rockmass stiffness, and other properties (Deere and Miller, 1966; Deere and
Deere, 1988; Cording and Deere, 1972; Coon and Merritt, 1970; Bieniawski,
1979). Philosophically, RQD provides a crude estimate of the percentage of the
rockmass which can be expected to behave in a fashion similar to a laboratory
sample (typically 10 cm long). A rockmass with a low RQD ( 50) has few intact
blocks larger than 10 cm. In such a rockmass, joints and fractures dominate the
rock's response to stress and gravity. The strength and stiffness of the rock, as
determined in the laboratory, has little relevance here. On the other hand,
rockmasses with RQD 95 % possess strength and stiffness much closer to the
values obtained in the lab. Joints may still dominate behaviour in low stress
environments but may have little or no influence at depth (provided joints are
clean and tight). Deere proposed the following categories of rockmass quality:
Table 2.14.2:
Rock Quality Designation ( Description) RQD Value
Very Poor 0 - 25
Poor 25 - 50
Fair 50 - 75
Good 75 - 90
Excellent 90 - 100
RQD does not accurately reflect conditions in rockmasses with joint spacings
greater than 0.3m. Rockmasses with block sizes in this range can be problematic
and therefore require additional parameters for adequate classification. While
RQD forms the starting point for most assessment procedures, more
comprehensive classification schemes are discussed in the following sections.
Also consider the directional nature of RQD. For example, a drill hole parallel
to a set of major laminations in highly anisotropic rock will yield a relatively high
RQD. Core taken from a hole perpendicular to the lamination set will give a much
lower value. When practical, drill core from two or more boreholes at different
angles should be considered for complete assessment of RQD.
RQD is intended to give a measure of in situ and undisturbed rockmass
conditions. Therefore all core breaks due to drilling, handling and discing must be
ignored in the calculation of RQD. In addition, minor cracks in the core which are
not related to established jointing should also be ignored. Failing to do so may
result in an overconservative or unrepresentatively low measure of RQD. In hard
rock mining applications, RQD typically measures between 50 and 100. Values
lower than this represent special conditions or an unusually poor rockmass.
Exceptions include RQD measured perpendicular to schistosity or foliation. Such
a measurement may be much lower than the RQD of the surrounding rock. Blast
damage to a rockmass can also be reflected in a reduced RQD (Løset, 1992).

Figure 2.14.4: Obtaining RQD from volumetric joint count, Jv
Ideally the RQD should be measured as soon as possible after the core has been
removed from the core barrel. RQD should be a part of the preliminary logging
procedure. If a clearly marked rule is laid out along side of the core box, a
geologist or technician can become very efficient at estimating RQD with a
minimum of additional time expenditure. The window size for RQD calculation
and recording can vary between 2 m and 10 m (e.g. record a separate measure for
each successive 2 m of core recovered) depending on the resolution required for
the project. The window should be reduced in zones of geological transition or
where the measured RQD is observed to change significantly over short distances.
In addition, RQD can be remeasured some time after recovery to determine if the
rock is susceptible to rapid disintegration upon exposure.
Alternate methods of estimating RQD
Unfortunately, the engineer is often required to estimate RQD without timely
access to drill core or without historical logs of RQD. Palmström (1982) proposed
a fallback method of estimating RQD from exposed joint traces on excavation
walls or outcrops. The Volumetric Joint Count, J , is the sum of the number ofV
joints per metre for each of the major joint sets present. Alternatively, the inverse
of the representative true spacings for each set can be used, as shown in Figure
2.14.4. Note that true spacings must be used and not the apparent spacings
produced by oblique intersection with the rock wall. This measure is valid for
rockmasses with 3 or more well developed joint sets. This estimated RQD will
represent a maximum value. That is, no random joints or fractures are considered
and damage due to blasting and stress are also ignored. Palmström (1995) gives
alternate relationships for one and two dominant joint sets, while Priest (1993) and
Priest and Hudson (1976, 1981) present RQD relationships using scanlines.
Figure 2.14.5: Estimating equivalent RQD ( RQD ) from exposed wall jointingW
Another simple method for estimating RQD is illustrated in Figure 2.14.5. A
two metre graded rule can be placed on an exposed rock face. Calculate RQD as
described for the drill core, considering any well developed joint which intersects
the ruler as a core break. When estimating RQD for an undisturbed rockmass, care
must be taken to consider only in situ joints and not induced tensile cracks and
blast related fractures. Disregard any fractures which are less than 30 cm in length
and consider disregarding larger fractures which are clearly induced (have a
sugary surface). This will give a best case or upper-bound value of RQD.
This technique can also be used to determine the degree of degradation due to
blasting and excavation. By considering all joints and fractures (induced) in the
measurement of RQD ( wall ), an estimate of post-excavation rockmass qualityW
can be obtained (ignore fractures less than 30 cm in length). This may be a more
relevant value for local support design. Note the additional subscript W attached
to this measurement. Maintain this notation to differentiate the value from true
RQD (joints only). Palmström's RQD (1982) and this crude RQD measurementW
serve to provide an upper and lower bound respectively for local rock quality.

2.14.4 Rock Mass Rating, RMR
This rating system is also known as the Geomechanics Classification and was
developed by Bieniawski (1976) for use in design of tunnels in hard and soft rock.
As more data was collected, Bieniawski continued to refine his classification
scheme making changes and adjustments as necessary. This discussion is based on
the latest revision detailed by Bieniawski (1989; 1993).
The scheme considers six factors: Factor Range
+ Uniaxial strength of the intact rock material: A1 0 - 15
+ Rock Quality Designation, RQD: A2 3 - 20
+ Spacing of discontinuities: A3 5 - 20
+ Condition of discontinuities: A4 0 - 30
+ Groundwater conditions: A5 0 - 15
+ Orientation of discontinuities
(adjustment for tunnels mines): B (-12) - 0
A numerical factor is assigned to each category above and the sum of these
factors yields the Rock Mass Rating, RMR:
RMR = A1 + A2 + A3 + A4 + A5 + B
Based on this relationship and the parameters which are described in detail in
the following pages, Bieniawski proposed the following rockmass classifications:
Table 2.14.3:
Rock Mass Class Description RMR
I Very Good Rock 81 - 100
II Good Rock 61 - 80
III Fair Rock 41 - 60
IV Poor Rock 21 - 40
V Very Poor Rock 0 - 21
NOTE: Bieniawski(1989) suggests that poor blasting can reduce RMR by up to 20%
The rockmass to be considered should initially be divided into geologically or
geomechanically distinct zones (e.g.; hangingwall granite, hangingwall schist, ore
zone, footwall gabbro, main fault zone, etc.). Each zone should be classified
separately. If during the classification process, significant changes in structural
character or in proposed excavation profile are noted, then additional subdivisions
and classifications should be made until all unique geomechanical zones are
identified and assigned a rating. These zones permit adaptation of design to local
conditions and provide an immediate reference for future planning.
Factor A1 - Strength of Intact Rock Material
Strength of the intact rock can be obtained from uniaxial compressive strength
tests (I.S.R.M., 1981) or from Point Load Index tests (Hoek and Brown, 1980):
Table 2.14.4:
Uniaxial Compressive
Strength (MPa)
Point Load Strength
Index (MPa)
Factor
A1
250 10 15
100 - 250 4 - 10 12
50 -100 2 - 4 7
25 - 50 1 - 2 4
5 - 25 n/a 2
1 - 5 n/a 1
Factor A2 - Rock Quality Designation, RQD
RQD is used in the RMR classification as a measure of structural integrity:
Table 2.14.5:
Rock Quality Description RQD
%
Factor
A2
Very Good Rock 90 - 100 20
Good Rock 75 - 90 17
Fair Rock 50 - 75 13
Poor Rock 25 - 50 8
Very Poor Rock 0 - 25 3
Factor A3 - Spacing of Discontinuities
Calculate the average true spacing for each joint set. Use the smallest of these
average values to determine Factor A3:
Table 2.14.6:
Minimum Average Discontinuity Spacing
( cm )
Factor
A3
200 20
60 - 200 15
20 - 60 10
6 - 20 8
6 5

Factor A4 - Condition of Discontinuities
Based on limited information about the character of the discontinuity surfaces
and using the broad categories listed below, Factor A4 can be estimated:
Table 2.14.7:
Description of Discontinuity Surfaces
(Roughness, Persistence, Separation, Wall Condition)
Factor
A4
Very Rough Surfaces, Not Continuous (Non-persistent),
No Separation (Full Wall Contact), Unweathered Joint Walls
30
Slightly Rough Surfaces, Moderately Persistent,
Separation 1mm, Slightly Weathered Joint Wall Rock
25
Slightly Rough Surfaces, Moderately Persistent,
Separation 1mm, Highly Weathered Joint Walls
20
Slickensided OR Gouge 5mm OR Separation 1 - 5 mm,
Continuous (Highly Persistent)
10
Soft Gouge (or Clay) 5mm Thick OR Separation 5mm
Continuous (Highly Persistent)
0
When more information is available and a higher degree of accuracy is
warranted (due in part by the dominant nature of this parameter with respect to
RMR), use the chart below to calculate Factor A4. Consider each subfactor
separately and then add up the results to obtain A4:
A4 = E1 + E2 + E3 + E4 + E5
Table 2.14.8:
Persistence
or Length
( E1 )
Separation
or Aperture
( E2 )
Surface
Roughness
( E 3 )
Infilling
or Gouge
( E4 )
Weathering
of Joint Wall
( E5 )
1 m
( 6 )
None
( 6 )
Very Rough
( 6 )
None
( 6 )
Unweathered
( 6 )
1 - 3 m
( 4 )
0.1 mm
( 5 )
Rough
( 5 )
Hard Infilling
5 mm
( 4 )
Slightly
Weathered
( 5 )
3 - 10 m
( 2 )
0.1 - 1.0 mm
( 4 )
Slightly Rough
( 3 )
Hard Infilling
5 mm
( 2 )
Moderately
Weathered
( 3 )
10 - 20 m
( 1 )
1 - 5 mm
( 1 )
Smooth
( 1 )
Soft Infilling
5 mm
( 2 )
Highly
Weathered
( 1 )
20 m
( 0 )
5 mm
( 0 )
Slickensided
( 0 )
Soft Infilling
5 mm
( 0 )
Decomposed
( 0 )
Factor A5 - Ground Water
Groundwater (or persistent mine water at depth) can play a significant role in
rockmass behaviour by altering joint surface conditions with time and by creating
an effective stress condition in which normal rock pressure is relieved across
joint surfaces, thereby reducing the frictional shear strength. RMR accounts for
this effect through the inclusion of Factor A5:
Table 2.14.9:
Inflow in litres per
10m tunnel length
Jnt. Water Pressure
Major Princ. Stress
General
Conditions
Factor
A5
None 0 Dry 15
10 0.1 Damp 10
10 - 25 0.1 - 0.2 Wet 7
25 - 125 0.2 - 0.5 Dripping 4
125 0.5 Flowing 0
Factor B - Joint Orientation Adjustment
The basic value of RMR, which is used for estimation of stiffness and strength
properties, and for classification of the rockmass independent of the proposed
excavation, includes the summation of the five A factors only. When RMR is to
be used for determination of support requirements and general stability
assessment, the relative orientation of dominant discontinuities with respect to the
excavation must be taken into account. In addition to adjustments for foundations
and slopes, Bieniawski (1989) provided the following classes for Factor B for
tunnelling and mining (note the negative adjustments to RMR):
Table 2.14.10:
Orientation of critical (most detrimental) joint set
with respect to tunnel or mine excavation
Factor
B
Very Favourable 0
Favourable -2
Fair -5
Unfavourable -10
Very Unfavourable -12
These descriptions (i.e. favourable, unfavourable, etc.) are based on the
consideration of relative strike and dip of the joint and of the excavation and on
the relative direction of tunnelling (development) with respect to joint dip as
summarized in Figure 2.14.6.

Figure 2.14.6: Effect of discontinuity orientation on stability - use with factor B
Factor B - Joint Orientation Adjustment (continued)
The orientation adjustments for joints striking perpendicular to the tunnel are
based on the assumption that patterned bolting is being installed with each round.
The difference between driving with dip and driving against dip arises from the
ability to safely bolt potential blocks, before they become liberated, when driving
with dip. When the discontinuity strike is neither parallel nor perpendicular to the
tunnel, use an adjustment based on dip which lies between these two extremes.
Additional Notes
Bieniawski (1989) summarizes a number of modifications which could lead to
improved applicability of the Geomechanics Classification, RMR, to mining. In
addition, Laubscher (1977, 1984), Laubscher and Taylor (1976) and Page and
Laubscher (1990) and Stacey and Page (1986) describe a classification system
based on RMR called the Modified Rock Mass Rating which accounts for blasting,
stress change, mining influences and reduced design stand-up times encountered
in mining.
Q
RQD
Jn
x
Jr
Ja
x
Jw
SRF
=
2.14.5 Rock Tunnelling Quality Index, Q
Barton et al. (1974) studied a large number of underground excavation case
histories and developed the Tunnelling Quality Index, Q. It is a means of
classifying rockmasses with respect to in situ parameters including rock quality,
joint condition and stress state. Q is defined by:
where:
RQD is the Rock Quality Designation (Section 2.14.3).
This parameter indicates the percentage of rock which can be expected
to possess strength and stiffness properties comparable to a 10cm
laboratory sample of intact rock. RQD ranges from 10 to 100 when being
used in the calculation of Q.
Jn is the Joint Set Number
This factor accounts for the number of repetitive joint sets and the
relative dominance of random fracturing and jointing. Jn ranges from a
value of 0.5 (no joints) to 20 (completely crushed rock).
RQD/Jn is a crude representation of the average block size.
The extreme values of RQD/Jn thus calculated are 0.5 to 200. Clearly this
is an extremely crude index of block size. It does, however, provide a
means of comparison and can be used to empirically estimate support
spacing and surface retention requirements.
Jr is the Joint Roughness Number
Jr describes the large and small scale surface texture of the critical joint
set. Jr ranges from 0.5 (unfavourable) to 4.0 (favourable).
Ja is the Joint Alteration Number
Ja describes the surface alteration and frictional resistance of the critical
joint set and ranges from 0.75 (favourable) to 20 (unfavourable).
Jr/Ja represents joint surface integrity and strength. It favours rough,
unaltered, discontinuous joints. When rocks are smooth or slickensided
(polished by shear) and/or if they contain low friction coatings or filling,
they are considered detrimental to stability. Jr/Ja for the critical joint set
should be used in the calculation of Q. This is the joint set most likely to
cause problems based on the values of Jr and Ja, and also based on the
geometry of the joint. Joints which make a shallow angle (35°) with
respect to a surface are the most critical, followed by joints parallel to a
surface. Joints which are perpendicular to an excavation surface are
usually the least critical. In the case where gravity sliding is the dominant
failure mode, inclined joints ( 35° with respect to horizontal) are likely
to be critical. Clearly, some subjective judgement is required here in
order to determine the critical set.

Q
RQD
Jn
x
Jr
Ja
x
Jw
SRF
=
Jw is the Joint Water Reduction Number
Jw accounts for the destabilizing effect of high water pressures and of
joint washout by water influx. Jw ranges in value from 1.0 for dry
excavations to 0.05 for excavations with excessive inflow and pressure.
SRF is the Stress Reduction Factor
SRF modifies Q to account for high in situ stresses which may cause
compressive failure of the rock. The destabilizing effect of very low
confining stresses in structured ground is also considered. Either
condition results in a higher value of SRF and therefore in a lower value
for Q. Optimum SRF (0.5) is achieved under moderate confining stress
which locks up the joint structure while posing no danger of overstress
of intact rock. SRF is also affected by large scale weakness zones which
may cause unfavourable conditions for excavation
Jw/SRF is a complex factor representing the active stress (and strength) state in
a rockmass as it occurs in situ and as it may be altered by the presence of
water and cross-cutting structural weaknesses.
Barton et al. (1974) proposed the following classifications of rockmass quality
based on the evaluation of Q:
Table 2.14.11:
Tunnelling Quality Index Rock Mass Description
0.001 - 0.01 Exceptionally Poor
0.01 - 0.1 Extremely Poor
0.1 - 1 Very Poor
1 - 4 Poor
4 - 10 Fair
10 - 40 Good
40 - 100 Very Good
100 - 400 Extremely Good
400 - 1000 Exceptionally Good
These classifications can be used to make relative comparisons between
different rockmasses. Q is used later in this chapter to determine stability of
excavations, to estimate strength and stiffness parameters and to make crude
support recommendations. The factors which make up Q are determined as shown
in the following tables. In hard rock mines, Q typically ranges from 0.1 to 100.
RQD Rock Quality Designation
RQD is calculated using drill core from the area of interest. Other methods of
calculating RQD in the absence of core samples are given in Section 2.14.3. The
fundamental interpretation of RQD is the same regardless of the method:
RQD = Sum of lengths of core sticks greater than 10cm long x 100
Total length of core run
Table 2.14.12:
Rock Quality Designation ( Description) RQD Value
Very Poor 0 - 25
Poor 25 - 50
Fair 50 - 75
Good 75 - 90
Excellent 90 - 100
NOTE: When RQD is less than 10, use 10 for the purposes of calculating Q and Q'.
Jn Joint Set Number
Jn is determined from the results of joint mapping, stereonet plotting and
cluster contouring as shown below. Note that a joint set must be relatively well
developed as a cluster. Otherwise it should be considered as a random joint.
Table 2.14.13:
# of Joint Sets Jn # of Joint Sets
Intact Rock
No Joints
0.5 1
Few Random
Joints Only
1 Set 2 3
1 Set
+ Random
2 Sets 4 6
2 Sets
+ Random
3 Sets 9 12
3 Sets
+ Random
4 Sets
Heavily Jointed
15 20
Earthlike,
Crushed Rock
NOTE: Stereonets should show local joints only (from current design zone ) for Jn

Jr Joint Roughness Number
Jr relates both large and small scale surface texture for discontinuities:
Table 2.14.14:
Large Scale: Planar Undulating Discontinuous
Small Scale:
Jr
(Critical
Set)
Slickensided 0.5 1.5 2.0
Smooth 1.0 2.0 3.0
Rough 1.5 3.0 4.0
Gouge-Filled
No Wall Contact
1.0 1.0 1.5
NOTE: Add 1.0 to Jr if mean spacing of critical joint set exceeds 3m
Ja Joint Alteration Number
Barton et al. (1974) offer a comprehensive listing of alteration classifications
and Ja factors. The following chart is abbreviated for hard rock mining:
Table 2.14.15:
Typical Description (Critical Joint Set) Ja
Tightly Healed 0.75
Surface Staining Only 1.0
Slightly Altered Joint Walls. Sparse Mineral Coating. 2.0-3.0
Low Friction Coating (Chlorite,Mica,Talc,Clay) 1 mm thick 3.0-6.0
Thin Gouge, Low Friction or Swelling Clay 1 - 5 mm thick 6.0-10.0
Thick Gouge, Low Friction or Swelling Clay 5 mm thick 10.0-20.0
Jw Joint Water Reduction
Jw accounts for the weakening effect of groundwater and for the effective
normal stress reduction due to water pressure. Consider mine water only if it is
persistent. Do not consider water inflow from temporary drilling, for example.
Table 2.14.16:
Joint Water Description
Pressure
( kPa )
Jw
Dry Excavation (Less than 5 litres/min locally) 100 1.0
Medium Inflow or Pressure 100-250 0.66
Large Inflow or High Pressure
No Joint Filling
250-1000 0.5
Large Inflow or High Pressure
Outwash of Joint Filling
250-1000 0.33
Exceptionally Large Inflow or Pressure
Decaying After Excavation
1000 0.2-0.1
Exceptionally Large Inflow or Pressure
No Reduction After Excavation
1000 0.1-0.05
SRF Stress Reduction Factor (a; rock stress)
SRF is used to account for fracturing of the rock due to overstressing during
excavation and to account for reduced confinement of structurally dominant
rockmasses near surface (or in late-stage, destressed mining environments).
Figure 2.14.7: SRF with respect to in situ stress. Note: for highly anisotropic stress:
when 5. /. 10, use . =0.8. ; when . /. 10, . =0.6.1 3 c c 1 3 c c

SRF Stress Reduction Factor (b; weakness zones)
SRF can also be used to account for major weakness zones in areas where their
presence dominates the behaviour of the rockmass and causes loosening when
excavated. Use these factors (SRF;b) instead of those on the previous page
(SRF;a) when these weakness zones influence or intersect the excavation:
Table 2.14.17:
Weakness Zone (If Present) SRF
Single shear zone in competent rock or
Single weakness zone containing clay,
and/or chemically disintegrated rock
Excavation Depth
50 m
50 m
2.5
5.0
Loose open joints, heavily jointed or 'sugar cube' (any depth) 5.0
Multiple shear zones in competent rock or
Loose surrounding rock (any depth)
7.5
Multiple occurrences of weakness zones containing clay,
and/or chemically disintegrated rock or
Very loose surrounding rock (any depth)
10.0
Swelling rock - chemical swelling due to presence of water 5-15
Squeezing rock - plastic flow of weak rock under stress 5-20
NOTE: Reduce these values by 25-50% when weakness zones influence but do not
intersect the excavation.
Additional Notes
+ For critical intersections and access portals use 2.0 to 3.0 x Jn to evaluate Q.
+ Where drill core is not available, see Section 2.14.3 to evaluate RQD.
+ Only well developed joint orientation clusters should be considered as sets.
Foliation sets should be considered only if the potential for significant parting
exists. Otherwise, in either case, consider the set as random.
+ When determining Jn, filter joint data set to include only those joints within a
reasonable spatial distance from the proposed excavation segment. All joint sets
may not be present in every location. Delineate design zones based on
convenient excavation steps (e.g. for each stope) or based on structural change
and obtain local values of Q.
+ The joint set with minimum Jr/Ja should be used to evaluate Q unless this joint
is favourably oriented for stability (non-sliding or perpendicular to clamping).
In this case, define a different critical joint. Jr/Ja relates to surfaces most likely
to initiate failure.
+ Use SRF;a if intact rock dominates stress response. Use SRF;b if the rock mass
contains clay or if large scale weakness zones are present. In this case, the
intact rock will play little role in stress response.
Q
RQD
Jn
Jr
Ja
Jw' = × ×
Q
RQD
Jn
Jr
Ja
' = ×
2.14.6 Modified Rock Quality Index Q'
The Rock Tunnelling Quality Index, Q has been used with a great deal of
success (Barton et al., 1992) in the design of tunnels in rock. It contains six
parameters which were deemed to influence the inherent stability of the rockmass
and which therefore dictate the degree of support required for tunnelling.
The parameter SRF, however, becomes redundant when the classification
system is used for the estimation of rockmass properties for the purpose of
analytical or numerical modelling for design. The influence of stress is taken into
account within the model.
The Modified Rock Quality Index Q' is given as;
SRF is set to 1.0 which is equivalent to a moderately clamped but not
overstressed rockmass. In addition, in most underground hard rock mining
environments, the excavations are relatively dry (not considering transient mine
water flow from drilling or backfilling). Jw, therefore can also be set to 1.0 in this
case. In environments with high water pressures, stress-based analyses should
include the effects of water pressures and flow if Jw is to be dropped from Q'.
When specifying the Modified Rock Quality Index, always use the apostrophe (')
to distinguish it from Barton's Q.
This parameter should better reflect the inherent character of the rock mass,
independent of the excavation size and shape which are considered separately in
subsequent analyses. Use Q' therefore to estimate rockmass modulus and strength
(Section 2.15 and Hoek et al., 1995) while using the original Q directly when
applying Barton's stability and support recommendations (Section 2.16).
Factor Q' with Jw also set to 1.0;
is used along with several other factors (accounting for jointing, stope geometry
and overstress) to determine the Modified Stability Number, N', which is used in
the Modified Stability Graph method (Mathews et al., 1981, Potvin, 1988;
Bawden, 1993 and Hoek et al., 1995) for dimensioning of open stopes in mining
and for the design of cablebolt support in these environments (Section 2.17).

Figure 2.14.8: Rockmass component contributions to RQD, Q, Q', and RMR
2.14.7 Comparison of Rockmass Classifications
The previous discussions of RQD, RMR, Q and Q' illustrate only a portion of
the development history of classification schemes. Classification methods for
tunnelling and rock excavation begin with an early reference by Terzaghi (1946),
followed by work by Lauffer (1958), Wickham et al. (1972), Laubscher (1977,
1984, 1993), Palmström (1995) and many others. In addition, other researchers
have modified RMR for application to mining problems (Laubscher and Taylor,
1976; Page and Laubscher, 1990; Cummings et al., 1982; Kendorski et al., 1983).
The treatment given in this book is merely intended as an initial and practical
introduction to the most popular methods. Interested readers are directed to the
preceding references. Comprehensive treatment of classification techniques can
be found with bibliographies in Hoek and Brown (1980), Hoek et al. (1995),
Franklin (1993) and Bieniawski (1989).
It is interesting to note the differences between the four systems presented here.
Figure 2.14.8 illustrates the relative emphasis in each system of specific
characteristics of the rockmass and of the environment. It is important to keep
these differences in mind when comparing and utilizing classification results.
Figure 2.14.9: Comparison between RMR and Q results (after Bieniawski, 1993)
In the initial stages of a project, it is always advisable to employ at least two
classification schemes and compare results and recommendations before
continuing (Kaiser et al., 1986). With increased experience at a particular site, it
may be adequate in time to maintain the use of only the most appropriate method
(i.e. that which seems to best predict behaviour as monitored).
Many relationships have been proposed which attempt to relate the outputs of
different rockmass classification schemes. In particular, the relationship
(Bieniawski 1979, 1993):
RMR = 9 ln Q + 44 or RMR = 21 log Q+ 44 or Q = 10e 10
(RMR - 44)/21
has been used throughout the literature. While this relationship can be useful for
comparison purposes or where correlations (e.g. rockmass modulus) are only
available for one classification system, it should not be used to apply the results
from one classification method to obtain recommendations based on another. This
caution is due to the differences summarized in Figure 2.14.8. Figure 2.14.9
below illustrates the scatter in the relationship and the hazards inherent in its use.

Figure 2.15.1: Influence of rockmass quality (as estimated by rockmass classification)
on in situ rockmass properties
2.15 Rockmass Properties
from Classification Systems
It is relatively simple to measure intact rock properties such as uniaxial
compressive strength and rock modulus (I.S.R.M., 1981) from laboratory
specimens. In situ fracturing (at all levels of scale), jointing, relaxation and
weathering all serve to degrade the properties of the rockmass. Measurement of
the real values of rockmass strength and stiffness is difficult and is beyond the
practical and economic scope of most mining operations.
One alternative is to use rockmass classification to adjust for in situ effects and
to estimate these parameters.
2.15.1 Rockmass Strength
A great deal of research (Deere, 1968; Hoek and Brown, 1980; Hoek, 1983;
Hoek and Brown, 1988; Hoek et al., 1995; and others) has focussed on the
estimation of triaxial strength characteristics of rockmasses. A detailed treatment
of these developments is beyond the scope of this chapter. In short, increased
fracture density has a dominant influence on the cohesive (confinement
independent) shear strength of rockmasses. Frictional (confinement dependent)
shear strength reduces as joint interlock is reduced, and internal (block) mobility
is increased. As a conservative estimate, the tensile strength of a heavily jointed
rockmass reduces to zero.
( ) { }
{ }
UCS s UCS s e GSI s
GSI
GSI RMR RMR GSI Q
cR L
GSI
e
= = = =
=
= − = +
−⎛
⎝
⎜ ⎞
⎠
⎟
σ where but for
Geological Strength Index (Hoek et al., 1995)
for or
100
9
25 0
5 23 9 44
;
ln '
σ σ σ σ σ
σ
σ
σ
1 3 3
2
1
3
F F
b C
F
C
F
F
C
m s
m s
= + +
=
=
=
=
=
major principal stress
at failure
minor principal stress
at failure
unconfined compressive
strength of rock
rockmass strength
parameters
m strength parameter for
intact rock
b
i
intact
,
SURFACECONDITION
VERYGOOD(Jr/Ja=3to5)
Veryrough,unweatheredsurfaces
GOOD(Jr/Ja=1to3)
Rough,slightlyweathered,ironstained
FAIR(Jr/Ja=0.3to1)
Smooth,moderatelyweatheredoraltered
POOR(Jr/Ja=0.05to0.3)
Slickensided,highlyweatheredwithcompact
coatings/fillingscontainingangularfragments
VERYPOOR(Jr/Ja0.05)
Slickensided,highlyweatheredwith
softclaycoatingsorfillings
Unconfined compressive strength, UCS , of a non-intact rockmass can beR
estimated from the laboratory strength UCS , and using either RMR (SectionL
2.14.4) or Q' (Section 2.14.6) using the following relationships:
Note, however, that even in ideal apparently intact rock, the uniaxial yield
strength (onset of damage) rarely exceeds 1/2 of the laboratory test value
(Bieniawski, 1976; Brace et al., 1966; Martin et al., 1993; 1994; 1995).
Hoek et al. (1995) present the summary in Table 2.15.1 for triaxial or
confinement dependent strength based on rockmass condition. Refer to this
reference for a more detailed discussion of rockmass strength.
Table 2.15.1: In situ rockmass strength and stiffness (after Hoek et al., 1995)
HOEK-BROWN CRITERION
ROCKMASS STRUCTURE
BLOCKY
(RQD/Jn 7.5)
Very well interlocked, undisturbed
rockmass consisting of cubical
blocks: 3 orthogonal joint sets
m /mb i
s
E(GPa)
v
GSI
0.60
0.19
75
0.20
85
0.40
0.62
40
0.2
75
0.26
0.015
20
0.25
62
0.16
0.003
9
0.25
48
0.08
0.0004
3
0.25
34
VERY BLOCKY
(RQD/Jn = 0.25 - 7.5)
Interlocked, partially disturbed
with polyhedral angular blocks
formed from 4 or more joint sets
m /mb i
s
E(GPa)
v
GSI
0.40
0.62
40
0.2
75
0.29
0.021
24
0.25
65
0.16
0.003
9
0.25
48
0.11
0.001
5
0.25
38
0.07
0
2.50
0.3
25
EXT. BLOCKY/SEAMY
(RQD/Jn 0.25)
Folded and faulted with many
intersecting discontinuities
forming angular blocks
m /mb i
s
E(GPa)
v
GSI
0.24
0.012
18
0.25
60
0.17
0.004
10
0.25
50
0.12
0.001
6
0.25
40
0.08
0
3
0.3
30
0.06
0
2
0.3
20

Figure 2.15.2: Influence of joint closure on rockmass modulus
2.15.2 Stiffness: Rockmass Modulus
The stiffness of a rockmass controls the response to loading and unloading at
levels of stress below the strength of the material. Elastic strains and therefore
displacements can be directly related to stress changes using the rockmass stiffness
or modulus. Rockmass modulus is required as input into most numerical and
analytical models of rockmass behaviour. Unfortunately, it is very difficult to
accurately estimate and, in mining, impractical to measure with any confidence.
The rockmass modulus, in simple terms, directly relates a uniaxial stress change
with an induced strain or a displacement increment in a unit length of rock. It is
a combination of the intact rock response and the combined responses of all of the
cross-cutting joints, as shown below.
As illustrated above, rockmass stiffness or modulus depends on the modulus of
the intact rock, on the joint density and on the joint surface character, wall
stiffness and infilling. The modulus of a fractured or jointed rockmass, however,
depends primarily on the level of confinement. Due to the lower unconfined
modulus, a 1 MPa stress increase (or decrease) at low confinement (e.g. near
surface) will result in much larger strains than will a similar stress increment at
higher confinements and depths. At lower confinements, higher strains result from
the requirements for initial closure to compress the softer joint surfaces or
infilling. The modulus significantly increases once full closure is achieved.
1 1 1
E E K SRM L N
= +
If K represents the closure stiffness of a joint (GPa/m) at a given range ofN
confinement, if E is the lab rock modulus (E in Figure 2.15.2) and if SL T50
represents the spacing of a single joint set perpendicular to the applied stress, then
the rockmass modulus, E is given by:RM
For anisotropic rockmasses, such as illustrated above, the stiffness obtained in
the above equation applies to the direction perpendicular to the joint surfaces. The
modulus parallel to the joints (e.g. parallel to an undamaged foliated hangingwall
with minimal cross-jointing) will be better represented by the intact rock modulus.
Models have been presented (Amadei and Goodman, 1981; Gerrard, 1982) to
calculate rockmass moduli (isotropic and directional) which incorporate multiple
joint sets. These models normally require measurement and input data which is
beyond the scope of mining operations and most construction projects.
Many large scale techniques have been developed to measure elastic modulus
and/or deformation modulus (ratio of applied stress to total of inelastic and elastic
strain) in the field (Bieniawski, 1978; Barton et al., 1980; Goodman, 1980;
Grimstad and Barton, 1993; Rocha and da Silva, 1970; Rocha, 1970; Serafim and
Pereira, 1983). These techniques include:
+ Radial Jacking Test (I.S.R.M., 1979)
+ Flat Jack Tests (I.S.R.M., 1986)
+ Goodman Jack (Goodman, 1972; Hustrulid, 1976)
+ Plate Bearing Tests (I.S.R.M., 1989)
+ Borehole Dilatometer (I.S.R.M., 1987; Hyett et al., 1992; Yow, 1993)
+ Large Scale Triaxial Tests (Müller, 1974; van Heerden, 1975)
+ Petite Seismique Technique (Heuzé, 1980; Bieniawski, 1979)
Back analysis techniques using modelling and instrumentation have also proven
useful for estimation of rockmass modulus:
+ Tunnel Relaxation (Waddell et al., 1970)
+ Pillar Monitoring (Wagner, 1974)
Rockmass classification schemes provide a practical alternative and facilitate
preliminary estimates of rockmass modulus ( ± 50 %). Correlations between Q,
RQD and RMR have been made using many of the techniques listed above. A
summary of this body of work is given on the following pages.

Figure 2.15.3: Rockmass modulus
from classification
Rockmass modulus from rockmass classification : Case histories
Many attempts have been made
by researchers and engineers to
relate rockmass classification
results to rockmass modulus as
measured by a wide variety of
field testing techniques.
RQD provides a measure of the
percentage of a rockmass volume
which can be expected to behave
in manner similar to a lab sample.
There is therefore a relationship
between RQD and the modulus
ratio; the ratio between the
modulus of the rockmass and that
of a standard lab sample. Note the
scatter, however, in this graph.
Barton et al. (1980) sought a
relationship between Q and
modulus. As the data is limited,
the scatter is great. Also note that
the evaluation of Q does not
involve the intact rock properties
even though the intact rock
modulus must govern at higher
values of Q. Q does incorporate a
measure of the clamping stress
which has a direct influence on
modulus of fractured rock.
RMR incorporates the
compressive strength of rock
which is related to modulus
(Deere, 1968). Two alternative
curve fits from different authors
are shown and seem valid for
RMR 50.
In all cases on this page the
applicability limits of the fitted
curves must be respected.
Figure 2.15.4: Modulus vs RQD
Figure 2.15.5: Rockmass Modulus vs Q' and RMR
Rockmass modulus from rockmass classification: Recommendations
The figure at right gives crude
limits for modulus-ratio estimation
using RQD. Note that higher stresses
tend to close fractures which in turn
increases the overall modulus. In
moderate to high stress environments
and in virgin ground, use the upper
design zone. In loose, destressed or
disturbed ground, use the lower zone.
The centre line represents an
expected relationship for a tight but
not overstressed rockmass. In
anisotropic rock RQD must be taken
in the direction of interest.
Below is a suggested range of absolute rockmass modulus with respect to Q'
1 and RMR 40. Note that Q' (RQD/Jn x Jr/Ja) is used here. RMR should not
include the joint orientation correction. In anisotropic rock, measure RQD and
spacing in the direction of interest. Use the design zones as shown to account for
the degree of stress and clamping in situ. The rockmass modulus is limited to a
maximum defined by the Young's Modulus (E ) of an undisturbed laboratoryT50
sample. Use the E directly for RMR 85 or Q' 100.T50

2.16 Empirical Design
Rockmasses can be extremely complex media for construction. It is often
difficult to apply mechanistic (based on physical mechanisms) analysis tools to the
design of excavations in rock. Rockmass classification methods have been
calibrated to provide alternative tools for this purpose. Classification and its
application to underground support design is primarily founded in civil
engineering tunnel construction. This is particularly true for RQD (Deere et al.,
1967), RMR (Bieniawski, 1976, 1989, 1993) and Q (Barton et al.,1974; Barton,
1988; Grimstad and Barton, 1993). Laubscher and Taylor (1976) and Laubscher
(1993) modified RMR for use in the design of block caving mines. In addition,
Mathews et al. (1981) and Potvin (1988) developed an extension of the Q system
and applied it to open stope design. Potvin's method is described separately in
Section 2.17 as it is specifically applicable to cablebolt design.
This section outlines some applications of RQD, RMR and Q to the
determination of unsupported excavation stability, stand-up time, general support
recommendations and specifically, to cablebolt design. It is important to
understand the origin of these empirical support methods. That is, they are
primarily based on tunnels at low to moderate depth (0 to 500 m). Civil tunnels
must be completely stable (no local block fallout) and must endure for many years
or decades. Recommendations for stability and support may therefore not be
directly applicable to mining. Wherever possible the authors of this handbook
have attempted to adapt the recommendations for mining openings and for
cablebolt support.
Cablebolt densities (number of cablebolts per unit face area) and bolt spacings
are primarily based on overall support pressure requirements (support load or
support capacity per unit face area) which have been derived for rockbolts. The
bolt densities prescribed for cablebolts are reduced by a factor corresponding to
the increased unit capacity of a cablebolt strand as compared to that of a rockbolt.
This increased spacing may allow local block fallout to occur between cables. It
is recommended that the reader refer to Stillborg (1986), Choquet (1991), Hoek
and Brown (1980), Hoek et al. (1995), or other rock support guides for
recommendations for rockbolting. Rockbolts and screen or straps should be used
in combination with the cablebolts to arrest this local unravelling where necessary.
In addition, the bolt lengths recommended by most empirical guidelines refer
to mechanically anchored rockbolts or resin grouted bolts. These devices have
fixed attachment points (anchor and head) or have highly adhesive bonds which
generate full load capacity over short anchor lengths ( 1 m). Cablebolts, however,
transfer their load to the rockmass over larger bond lengths. In addition, the top
of a cablebolt hole may contain a void of 0.5 to 1 m depending on the installation
method and quality control. For this reason a minimum anchor length of 2 m has
been added to all length guidelines to adjust them for cablebolting applications.
Figure 2.16.1: RQD-based stability and support guidelines (after Merritt, 1972)
2.16.1 Rock Quality Designation, RQD
Deere et al. (1967, 1969) have developed tunnel support guidelines for
different excavation methods based on RQD. These recommendations have
limitations due to the limited scope of RQD as a rockmass quality indicator.
Clearly, the influence of joint condition, rock strength and confinement (field
stress) are ignored in the calculation of RQD. As a crude measure of structural
integrity, RQD can be a convenient tool for preliminary design. For tunnel spans
of 6 to 12 m, Deere et al. (1969) proposed the approximate relationship:
Bolt Spacing (m) for a Square Pattern = 0.02 × RQD (%)
This relationship implies that cablebolts as primary support of tunnels 5 m
wide should be economically feasible for RQD values greater than 70% (Fair to
Excellent Rock) if 1.4 m is taken as the practical minimum cable spacing. A tight
pattern of mechanically anchored bolts and mesh are recommended below this
value. Shotcrete becomes a competitive support option below RQD = 60%.

Figure 2.16.2: Unsupported Tunnel Limits (after Bieniawski, 1993, 1989).
2.16.2 Rock Mass Rating, RMR
No-Support Limits and Stand-up Time
The Rock Mass Rating, RMR was originally developed by Bieniawski (1973)
and updated in 1979. Other authors have modified RMR for specific applications:
Mining:
Laubscher (1977, 1993); Kendorski et al. (1983)
Coal Mining:
Ghose and Raju (1981); Newman (1981); Sheorey (1993); Unal (1983);
Venkateswarlu (1986)
Slope Stability:
Romana (1985, 1993)
In Figure 2.16.2, Bieniawski (1993) presents the revised chart relating Span
and Stand-up time with his 1989 Rock Mass Rating System. The points in this
graph represent groundfalls in tunnels and in mining excavations. The concept of
stand-up time was originally conceived by Lauffer (1958, 1960), to represent the
duration of time within which an excavation will remain serviceable and after
which significant instability and caving occurs.
Figure 2.16.3: Alternative representation of Figure 2.16.2 stand-up time guidelines
In order to use Figure 2.16.2, first determine the RMR for the rockmass in
question. The intersection of a specified RMR contour with the bottom of the
shaded zone gives the maximum span which can remain stable indefinitely without
support. Within the shaded zone, the RMR contour line gives the anticipated stand-
up time without support. Above the shaded zone (e.g. a 20 m span with RMR=60)
unsupported excavations will disintegrate shortly after development. Note the
range of data for which this relationship was derived.
In order to present these guidelines in a manner consistent with other systems,
Figure 2.16.2 has been replotted with RMR on the horizontal axis, as shown in
Figure 2.16.3. For a temporary mining opening such as a 10 m topsill (e.g. with
a required stand-up time of 1-2 months) it can be seen that a rockmass with a Rock
Mass Rating of greater than 65 may not need support (apply an appropriate safety
factor - multiplier of 2) with the exception of pinned screen for personal safety.
Note that poor blasting can reduce RMR by up to 20% (Bieniawski, 1989).
Following logic developed by Barton et al. (1974) and Barton (1994) for the Q
system, RMR can be increased by up to 10% (RMR 30%) for near vertical stope
walls. Note that the full RMR including joint orientation adjustment is used here.

RMR - Support Guidelines
Bieniawski (1979, 1993) presents support guidelines for a 10 m wide
horseshoe shaped, drill and blast tunnel under 25 MPa of vertical stress:
Table 2.16.1: Support recommendations from RMR (after Bieniawski 1993)
RMR Excavation
(Horseshoe,
10m Span)
Combined Permanent Support
20mm Rockbolts
Fully Grouted
Shotcrete Steel
Sets
81-100 Full face, 3m
advance.
None (Spot
Bolting if req'd).
None None
61-80 Full face, 1.5m
advance.
Install support
20m from face.
Bolts in crown 3m
long,
2.5m spacing,
Some mesh.
50mm as
required in
crown.
None
41-60 Top heading
bench, 1.5-3m
advance in top
heading with
rapid support.
Full support
10m from face.
Systematic bolts
4m long, 1.5-2m
spacing in walls
and crown. Mesh
in crown.
50-100mm
in crown and
100mm in
sides.
None
21-40 Top heading
bench, 1.0-1.5m
advance in top
heading with
immediate
support. Full
support within
10m of face.
Systematic bolts
4-5m long, spaced
1-1.5m in crown
and walls with wire
mesh.
100-150mm
in crown and
100mm in
sides.
Light to
medium ribs
spaced 1.5m
where required.
20 Multiple drifts
0.5-1.5m
advance in
headings. Install
full support
immediately.
Systematic bolts
5-6m long, 1-1.5m
spacing in crown
and walls with wire
mesh. Bolt invert.
150-200mm
in crown.
150mm in
sides and
50mm on
face.
Medium to
heavy ribs
spaced 0.75m
with lagging
and forepoling.
Close Invert.
This excavation could be loosely equated to a typical mining haulageway at
moderate depth up to 1000m, although the support recommendations will be very
conservative for such an application.
Note that the bolt spacings and shotcrete thicknesses, etc. in this table are
specified for a combination support system as listed. Do not extract, for example,
bolt spacings from this table for use as a single component system as this will
result in under-designed support.
RMR - Semi-Empirical Support Guidelines
RMR has been adapted by various authors for support design. Figure 2.16.4
illustrates one such development by Unal (1983). The concept is simple and yet
it produces reasonable results for cablebolt length and moderately conservative
(for mining applications) recommendations for cablebolt density. For extremely
poor rocks (RMR 10), the height of the zone requiring support is assumed to be
equal to the span. This height is modified by RMR as shown until the rock is
completely self-supporting at RMR=100%. Support pressure is the amount of
distributed load applied to the surface of the excavation (roof in this case) to resist
further displacement of the rockmass. It is assumed that the cablebolts used here
are pre-installed or installed at the excavation heading immediately after blasting
and that stiff systems are in use (modified strand or well grouted and plated plain
strand cablebolts).
The cablebolt density as plotted here refers to the quantity of complete
cablebolts (single or double strand cablebolts) per square meter of excavation face
area. This is a convenient measure of cable distribution since it gives support
pressure directly when multiplied by the capacity or tension in the steel cable.
Cablebolt length is simply calculated as the height of the supported rock zone with
an extra 2 m added to provide a minimum reliable anchor for the fully grouted
cablebolt. Cablebolts typically have at least 0.5 m of poorly placed grout at the
top of an uphole. This inactive length increases when quality control is poor or
when a quantity of grout flows away into fractures after placement, reducing the
grouted column height. If either condition is suspected, increase this anchor length
accordingly. 2 m represents the minimum prudent anchor design.
The cablebolt densities in Figure 2.16.4 are calculated for a rock specific
weight of 26 kN/m and for steel capacities of 20 tonnes (200 kN) for single strand3
and 40 tonnes for double strand cablebolts. These values correspond to the onset
of inelastic yield and should be used for permanent installations. For temporary
and non-critical openings, 25 and 50 tonnes can be used respectively,
corresponding to ultimate breaking strength of cables. This results in a 20%
decrease in cablebolt density as noted in Figure 2.16.4.
Stimpson (1989) further developed this concept of a supported height,
incorporating the influence of in situ stress ratio and excavation height:span ratio.
The shape of the loosening zone becomes an ellipse with its long axis oriented in
the direction of major principle stress. this is due to the confining effect of higher
stresses. Note that the opposite trend will be observed if the stresses are high
enough to cause rockmass failure. In this case the ellipse will be oriented with the
long axis perpendicular to the major principle stress. The modified loading height
is then modified as a function of RMR. Detournay and St. John (1988) present a
method for calculating the depth of failure around deep circular openings in
anisotropic stress fields. RMR can be used to obtain rock strengths for this model.

Figure 2.16.4: Tunnel support pressure, cablebolt length and density guidelines with
respect to span and RMR (based on Unal, 1983)
2.16.3 Rock Tunnelling Quality Index - Q
No Support Limits
Barton et al. (1974), Barton (1988, 1994) describe the application of the Q-
system for rockmass classification to the determination of no-support limits for
various types of excavations. Approximately 200 case examples were originally
classified to originally calibrate this system. Since then over two thousand new
empirical tunnel and large cavern designs have been successfully carried out
(Barton et al., 1992). Figure 2.16.5 shows the original database of supported and
unsupported excavations. The shaded zone represents the limits of practical
support application. The lower boundary of this zone is the limit of stability for
unsupported excavations of a given Equivalent Span, ES = Span/ESR, where:
Table 2.16.2:
Type of (after Barton, 1988)
Excavation
Number
of Cases
ESR
Temporary mine openings. 2 approx.
3-5 ?
Permanent mine openings; Low pressure water tunnels;
Pilot tunnels; Drifts and headings for large openings.
83 1.6
Storage caverns; Water treatment plants; Minor road and
railway tunnels; Surge chambers; Access tunnels, etc.
25 1.3
Power stations; Major road and railway tunnels; Civil
defense chambers; Portals; Intersections.
79 1
Underground nuclear power stations; Railway stations;
Sports and public facilities; Factories.
2 approx.
0.8 ?
Excavation Support Ratio, ESR is a factor used by Barton to account for
different degrees of allowable instability based on excavation service life and
usage. Divide the span of the excavation by the appropriate ESR value to obtain
the equivalent span for use in Figures 2.16.5 and 2.16.7.
Note that the number of mining case histories leading to the recommendation
of ESR = 3 to 5 for temporary mine openings is limited. Based on the authors'
experience, a maximum of 3 is recommended for mine openings unless local
experience justifies an increase.
Certain mining excavations are more critical than others from both an
operational and a safety point of view. Figure 2.16.6 provides no-support limits
in order of decreasing reliability, relating them to Barton's original ESR values.
Figure 2.16.6 is plotted against actual excavation span.

Figure 2.16.5: Case history database for Q-System (after Barton, 1988)
Figure 2.16.6: Q-system; No-support span limits for underground mine openings
Figure 2.16.7: Tunnelling Support Guidelines (after Grimstad et al., 1993).
Bolt lengths have been modified for cablebolting.
Q - Support Guidelines
Support recommendations based on the Q-system have evolved over the years
as more and more case histories have been added to the database. Barton (1988)
presented a tabulated series of detailed support recommendations based on
different combinations of rock quality, Q, and on Equivalent Span (Span/ESR).
Grimstad et al. (1993) proposed a summary graph based on these
recommendations which is designed to accommodate advances in shotcrete
technology. A version of this graph is shown in Figure 2.16.7. Again, this graph
was developed for permanent support in civil tunnels, shafts and caverns. These
recommendations are likely to be too conservative for mining. Cable lengths
shown on the right side are valid for ESR = 1. For greater values of ESR, these
lengths should be increased in accordance with actual span. A reasonable rule-of-
thumb for mining would be to multiply the lengths shown by (ESR) .0.5
Barton et al. (1974) recommend the following adjustments to Q for vertical
walls (Qw) to account for the reduced demand for support on the wall:
Q = 5×Q for Q10, Q = 2.5×Q for 0.1Q10 and Q = Q for Q0.1w w w
Caution should be used when combining the above adjustments with large
values of ESR ( 2 ). It is possible that unconservative designs may result.

p
Q Jn
Jr
=
−
.2
3
1
3
( )
p
Q
Jr
=
−
.2
1
3
Figure 2.16.8: Cablebolt spacings for mining excavations (Jn 6)
Support Pressure and Bolt Spacing
Barton et al. (1974) proposed relationships for support pressure, p, in MPa;
For more than 2 joint sets (Jn 6); For 0 - 2 joint sets (Jn 6);
If grouted cablebolts are installed prior to or immediately upon excavation of
the face in question, this support pressure can be crudely related to the installed
cablebolt capacity per unit area of excavated rock face or to an equivalent bolt
spacing. This relationship is plotted in Figure 2.16.8, for Jn6, using the initial
yield-strength of steel cable (200 kN/strand). A spacing increase of 10% as noted
implies the use of the short term breaking-strength (250 kN) and can be used in
temporary and non-critical applications. Note that this relationship does not
consider excavation span. This implies that a surface reinforcement action is
involved creating a self-supporting rock span. The cablebolt must limit internal
displacements and therefore must be stiff (modified geometries). In addition, in
the case of double-strand bolts and wider spacings, plates and intermediate rock
bolting may be required to maintain surface integrity between cables.
Figure 2.16.9: Limits of cablebolt application ( base graph after Hoek, 1981)
2.16.4 Empirical Cablebolt Design - General Limits
Cablebolt support is ideally suited for moderate to large openings in blocky
ground under low to moderate stress. This design zone is shown in Figure 2.16.9
with respect to RMR, Q, induced stress and rock strength. Cablebolt effectiveness
is limited to the following broad conditions:
+ Highly fractured and soft rockmasses at low stress levels will tend to unravel
between cablebolts spaced within economic limits. More continuous forms of
restraint (e.g. mesh) and reinforcement (e.g. shotcrete, rebar) are required as
surface retention to make the cablebolt support more effective.
+ Highly fractured rocks at elevated stress levels may exhibit squeezing and
disintegration which cannot be arrested effectively by cablebolts. Plating is
necessary along with intermediate bolting, mesh and/or yielding linings.
+ Blocky or Massive rocks at high stress levels are subject to unpreventable
spalling and ultimately to violent rupture which cannot be prevented by cables
alone. Cables, in combination with other restraint and reinforcement elements
can preserve the integrity of the broken rockmass after such brittle failure
depending on the severity of the overstress.

Figure 2.16.10: Bolt lengths in current practice (after Lang and Bischoff, 1984) with
adjustment for cablebolt application (relationships are for S.I. units)
2.16.5 Empirical Design - Rules of Thumb
Classification systems serve to differentiate between different rockmasses and
to adjust design accordingly. Rules of thumb for support design have been
developed for blocky to fractured ground (U.S.C.E. 1980; Lang, 1961; Farmer and
Shelton, 1980; Coates and Cochrane, 1970; Laubscher, 1984). These are based on
tunnels, caverns and mine openings and summarize current practice.
Most of these guidelines are designed for rockbolting (mechanical or resin
grouted) and as such can be used to select spacings for face support to supplement
cablebolting in fractured ground. In many cases the recommended spacings will
not be economically practical for use directly with cablebolts.
The lengths quoted in these rules of thumb should be adjusted for cables by
adding a minimum of two (2) extra metres of embedded length (unless it is
indicated that this adjustment has already been made by the authors as is the case
in Figure 2.16.10). Extrapolating to obtain cable lengths for spans greater than
those shown in these figures is not recommended. The figure boundaries represent
the applicability limits based on the source data. Figure 2.16.10 illustrates a data
set of rockbolt lengths in existing tunnels and caverns.
The U.S. Army Corps of Engineers (1980) developed a suite of simplified
recommendations for rockbolt spacing, length and support pressure, summarized
in Table 2.16.3. Rock loads are based on support pressure from actively tensioned
mechanical bolts and may be inappropriate for cablebolting. These guidelines, like
all others in Section 2.16 should be used in conjunction with other design tools.
Table 2.16.3: Bolting Guidelines (after U.S.C.E., 1980)
Maximum Spacing Recommendations (U.S.C.E., 1980):
Least of:
+ 0.5 times the bolt length
+ 1.5 times the width of critical and potentially unstable rock blocks
+ 1.8 m (applies only to rockbolts where screen is to be attached)
Minimum Spacing (U.S.C.E., 1980)
0.9 to 1.2 m (for cablebolts, 1.4 m is normally the economic limit)
Minimum Average Expected Bolt Loads (Applies Only to Stiff Cable Systems)
Roofs / Backs: Equivalent to weight of slab with thickness 0.2 times the span
Sidewalls: Equivalent to weight of slab with thickness 0.1 times the height
Minimum Bolt Length

Farmer and Shelton (1980) collected case histories from numerous authors and
formulated the recommendations for rock bolting in Table 2.16.4. Of particular
interest are the comments regarding support function and design philosophy.
Table 2.16.4: (after Farmer and Shelton, 1980)
Span
( m )
Number
( Dip- ) of
Joint Sets
Bolt Recommendations Comments
(after Farmer and Shelton by
authors of this handbook)
15 1 to 2
(0 to 45)
Length = 0.3 x Span
Spacing 0.5 x Length
Install bolts perpendicular to
lamination with mesh to
prevent flaking. Decrease
spacing in weak strata.
Bolting creates load carrying
beam over span. Grouted bolts
or modified cable strand create
higher joint shear stiffness.
Tension bolts (plate cables) in
weak rock. Angle bolts where
joints are vertical.
15 1 to 2
(45 to 90)
For wall bolts:
Installed at 90- to lamination
Length Height x cos (Dip)
Installed Horizontally
Length Height / tan (Dip)
(Dip = dip of joints)
Roof bolting as above. Side
bolts designed to prevent
sliding along planar joints.
Spacing should be such that
bolt capacity is greater than
sliding or toppling weight.
Tension bolts (plate cables).
15 2 with
tight
clean
surfaces
Length 2 x Spacing
Spacing 3 to 4 x Block Size
Install bolts perpendicular to
lamination with mesh to
prevent flaking. Decrease
spacing in weak strata.
Bolts should be installed
quickly after excavation to
prevent loosening and retain
tangential stresses. Tension
and plate to improve radial
confinement. Sidewall bolting
where wedge toes daylight into
excavation.
15 2 Alternate Primary (1) and
Secondary (2) Bolting:
Length(1) 0.3 x Span
Spacing(1) 0.5 x Length(1)
Length(2) 0.3 x Spacing(1)
Mesh to prevent spalling
Primary bolting supports span
and major blocks. Secondary
bolting retains surface blocks.
Limit spacings (and provide
load capacity) accordingly.
Bolt or cable lengths should
penetrate beyond extent of
known discrete wedges.
15 2 with
tight
clean
surfaces
Alternate Primary (1) and
Secondary (2) Bolting:
Length (1) 0.3 x Span
Spacing(2) 3 to 4 x Blk. Size
Length(2) 2 x Spacing(2)
Mesh as required for surface
block retention
2.17 Empirical Design of Open Stopes and Support:
Mathews/Potvin Stability Graph Method
Classical empirical tools such as RMR and Q were developed from a database
composed primarily of civil engineering tunnels at low to moderate depth. These
tools have proven invaluable to the tunnelling engineer. The recommendations
derived from these systems for dimensioning and support, however, often result
in conservative designs for large temporary or non-entry mining excavations.
While these systems are appropriate for high traffic mining roadways,
lunchrooms and equipment rooms where stability must be paramount, they are
difficult to apply to the problem of dimensioning and support design for large
open stopes. These limited access areas can be designed as temporary structures
and in the case of non-entry stopes, can tolerate limited local fallout of small rock
blocks provided that dilution is minimized and overall stability is maintained.
These criteria permit a more economical design suitable to mining.
RMR (Bieniawski, 1989; 1993) allows for design modification based on
reduced stand-up times for mining while Q (Barton et al., 1974) attempts to
include mining applications through the use of Equivalent Support Ratio.
Laubscher and Taylor (1796) modified RMR and introduced a classification
system for caving operations and for stability of mining excavations. Readers are
referred to Hoek et al. (1995) for additional discussion of these methods.
Large scale open stoping methods such as Vertical Crater Retreat, AVOCA,
Longhole and Blasthole Stoping rely on the selection of a limiting stope
dimension. Ideally these stopes can be designed to be self supporting. When
ground conditions or the need for larger stopes mandates the use of support,
cablebolting is the most logical choice and has been successfully applied. Mathews
et al. (1981) proposed an empirical method for the dimensioning of open stopes
based on Q' and on three factors accounting for stress, structural orientation and
for gravity effects. The method is used to dimension each face of the stope
separately based on a combination of these three factors and on the hydraulic
radius (calculated as surface area / perimeter ) of the face. The hydraulic radius
accounts for shape as well as size of the face.
Potvin (1988) modified this original method and calibrated it using 175 case
histories. Nickson (1992) added case histories and further investigated Potvin's
support design guidelines. These case histories include hangingwalls, footwalls,
ends and backs from a wide variety of mining environments. Other case histories
can be found throughout recent literature (Bawden, 1993; Bawden et al. 1989;
Greer, 1989). The method has been expanded by the authors in this handbook to
provide improved support guidelines.

Q
RQD
Jn
x
Jr
Ja
' =
2.17.1 Modified Stability Number, N'
The classification of the rockmass and of the excavation problem itself is
accomplished in the Modified Stability Graph Method through the use of the
Modified Stability Number, N', as specified by Potvin (1988), Potvin and Milne
(1992) and Bawden (1993). This parameter is similar to the value N proposed by
Mathews et al. (1981) but has different factor weightings. Canadian mines use
Potvin's N' while at present mines in Australia, for example, use Mathews' analysis
and N. Only N' (Potvin) will be considered here. This method has been referred
to as the Potvin method, the Mathews/Potvin method, the Modified Stability Graph
method and the Stability Graph method. The latter label will be used for the rest
of this discussion for clarity and brevity.
N' is based initially on Q', where;
and where;
RQD/Jn is a measure of block size for a jointed rock mass
Jr/Jn is a measure of joint surface strength and stiffness
Modified Stability Number N';
N' = Q' x A x B x C
where;
A is a measure of the ratio of intact rock strength to induced stress.
As the maximum compressive stress acting parallel to a free stope
face approaches the uniaxial strength of the rock, factor A degrades
to reflect the related instability due to rock yield.
B is a measure of the relative orientation of dominant jointing with
respect to the excavation surface. Joints which form a shallow
oblique angle (10-30-) with the free face are most likely to become
unstable (i.e. to slip or separate). Joints which are perpendicular to
the face are assumed to have the least influence on stability.
C is a measure of the influence of gravity on the stability of the face
being considered. Overhanging stope faces (backs) or structural
weaknesses which are oriented unfavourably with respect to gravity
sliding have a maximum detrimental influence on stability.
Table 2.17.1: Range of values (*for hard rock mining):
Range RQD/Jn Jr/Ja A B C N'
Maximum 0.5 - 200 0.025 - 5 0.1 - 1 0.2 - 1 2 - 8 0.0005 - 8000
Typical* 2.5 - 25 0.1 - 5 0.1 - 1 0.2 - 1 2 - 8 0.1-1000
Figure 2.17.1: Rock Stress Factor A (Potvin, 1988) for Stability Graph analysis
2.17.2 Stability Graph Method - Input Parameters
Compute the value of hydraulic radius, HR:
+ HR = Area (m ) = w x h (units of m)2
Perimeter (m) 2(w + h)
where A and B are the two dimensions defining the stope face to be analyzed.
Compute the modified stability number, N':
+ Measure or calculate the value of RQD, Jn, Jr and Ja
as described in Section 2.14.5
+ Compute Q'=RQD/Jn x Jr/Ja.
From the charts that follow:
+ Evaluate Rock Stress Factor A.
+ Evaluate Joint Orientation Factor B
+ Evaluate Gravity Adjustment Factor C
+ Obtain N'=Q' x A x B x C
Plot point (HR,N') on stability graph and determine stability and design zone.
Rock Stress Factor A

Figure 2.17.2: Determination of Joint Orientation Factor, B,
for Stability Graph analysis
Joint Orientation Factor, B
Figure 2.17.3: Estimation of true interplane
angle and Joint Factor B
Joint Orientation Factor, B:
Example Determination
The true angle between two planes
is not immediately given by the
relative dips and strikes of the planes.
It must be calculated as shown on the
following page or estimated from a
stereonet as in this example.
Consider the hangingwall face and
associated joint sets (Figure 2.17.3a).
Determination of B involves only the
pole to the face and the mean poles
for each joint set 1, 2 and 3.
Using a series of small circles
(cones) centred on the face pole, the
angle (cone angle) from this pole to
each of the joint set poles can be
estimated as in Figure 2.17.3b).
These small circles (cones) can be
generated by hand (Goodman, 1980;
Priest, 1985) or they may be
automatically generated by a
computer program such as DIPS
(Hoek et al., 1995) as shown here.
Cones drawn at 10, 30, 45, 60, and
90 degrees provide sufficient
resolution to determine factor B.
The true angle between planes is
given by the smallest angle between
poles to the planes. Figure 2.17.3.b)
illustrates how to determine that the
angle from the face to set 1 = 20-,to
set 2 = 53-, and to set 3 = 71-.
In Figure 2.17.3c), the angle
contours have been replaced by
corresponding Joint Orientation
Factors ( B ). This shows clearly that
joint set A is critical and that the
factor, B, should be set to 0.2 for the
Stability Graph analysis.

Joint Orientation Factor, B:
Direct Calculation of Interplane Angle
It is possible to determine directly the true interplane angle between the stope
face (wall plane) and the joint plane using the following simple procedure.
Given the Dip and the Dip Direction for a plane, the Trend and Plunge of the
corresponding pole (normal vector) can be calculated:
T = Trend = Dip Direction +180-
P = Plunge = 90- - Dip
For a stope wall plane, w, and a joint plane, j, the direction cosines with respect
to the global coordinate grid ( North, East, Down ) are denoted by N, E and D
respectively and are calculated as follows:
For the stope wall: For the joint plane:
N = cos( T ) ( cos( P ) N = cos( T ) ( cos( P )w w w j j j
E = sin( T ) ( cos( P ) E = sin( T ) ( cos( P )w w w j j j
D = sin( P ) D = sin( P )w w j j
Next calculate the dot product, w((j, between the wall face and the joint plane:
w((j = N N + E E + D Dw j w j w j
Finally, the true interplane angle, , is given by:
= cos ( w((j ) = acos( w((j )-1
This calculation can easily be solved using a calculator or can be implemented
in a spreadsheet or computer program.
Once this true interplane angle is calculated, it is possible to assign a Joint
Orientation Factor, B.
Figure 2.17.4: Simplified special cases for
determining factor B
Joint Orientation Factor, B
Simplified approach (special cases)
It is important to remember that measurements such as Dip and Dip Direction or
Strike are made relative to a global coordinate system. They cannot be used
directly to calculate the true angle between two planes since the applicable
coordinate system must be changed to be relative to one of the faces. Therefore the
procedures discussed on the previous pages must be implemented.
The calculation of interplane angle is simplified, however, when one of the planes
is approximately horizontal or near vertical (Dip . 0 or Dip . 90). In the case of
true angle calculation for determination of Factor, B, this condition must apply to
either the stope face or the joint plane (or both).
Horizontal Joint or
Horizontal Stope Face
(Back):
Consider only the difference in
Dip between the stope face and
the joint plane using the graph
at right to determine B. When
one plane is approximately
horizontal, then the difference
in Dip approximates the true
interplane angle.
Near Vertical Joint or
Near Vertical Stope Face:
The difference in Strike (or in
Dip Direction) must also be
considered in the case of
vertical features. Note that this
relationship as presented by
Potvin (1988) should only be
used when one of the planes is
near vertical.

Figure 2.17.5: Determination of Gravity Adjustment Factor, C,
for Stability Graph analysis
Gravity Adjustment Factor, C
Figure 2.17.6: Hydraulic Radius, HR
Hydraulic Radius
Before proceeding with the
application of the Stability Graph, it
is necessary to understand the nature
of the hydraulic radius, HR. In short,
HR is calculated by dividing the area
of a stope face by the perimeter of
that face as shown at right.
Most classification systems (e.g.
RMR and Q) define stability and
support zones with respect to a single
value of span. This is because these
methods are derived from tunnelling
databases in which the long span can
be assumed to be infinite and in
which the short span is therefore the
critical dimension. If this short span
is kept constant and if the long span
is reduced (to square dimensions, for
example), the stability increases as a
result of the increased confinement
and rigidity provided by the extra
two abutments. A face with a
dimension ratio greater than 10:1 can
be treated as a (tunnel) span
equivalent to the shorter dimension.
Hydraulic radius more accurately
accounts for the combined influence
of size and shape on excavation
stability. It is useful to become
familiar with the range of spans for
a given hydraulic radius. This will
provide a means of comparison with
other design methods which do not
use hydraulic radius. Figure 2.17.6
illustrates these limits for a fixed
hydraulic radius of 5 m. Note that
although it is possible to apply this
method to mining tunnels, the
method has been calibrated for open
stopes with finite dimensions and
with lower priority for safety.

Figure 2.17.7: Database (Potvin, 1988; Nickson, 1992) of unsupported stopes
2.17.3 Open Stope Case History Database
No-Support Limit
176 case histories by Potvin (1988) and 13 by Nickson (1992) of unsupported
open stopes are plotted on the Stability Graph shown below. The modified
stability number, N', and the hydraulic radius, HR, were calculated for each case
study as outlined in the previous sections. Stable stopes exhibited little or no
deterioration during their service life. Unstable stopes exhibited limited wall
failure and/or block fallout involving less than 30% of the face area. Caved stopes
suffered unacceptable failure. Potvin plotted a Transition Zone defined by these
cases to separate the Stable zone from the Caving zone. The upper boundary of
this zone represents a recommended no-support limit for design. For a calculated
value of N', determine the maximum hydraulic radius for a stable stope face.
Figure 2.17.8: Database of cablebolt-supported stopes
Limits of Cablebolt Effectiveness
Potvin (1988) and Potvin and Milne (1992) also collected 66 case histories of
open stopes in which cablebolt support had been used. Nickson (1992) added an
additional 46 case studies to this database which is illustrated below. Cablebolted
stopes exhibit improved stability leading to larger stable spans (greater hydraulic
radii). While this database does not take into account issues such as quality
control, it does provide a reasonable demonstration of cablebolt effectiveness.
Potvin plotted a limit for cablebolt effectiveness which Nickson modified using
statistical methods and additional data. The upper curve plotted below represents
the limit of reliable cablebolt performance. Nickson proposed a zone as shown
below to indicate the maximum stable hydraulic radius for cablebolted stopes
(upper bounding curve) and the reduction in confidence until cables can no longer
be assumed to be providing any degree of useful stope support (lower bound).
Below this zone caving is inevitable.

Figure 2.17.9: Design Zones for Open Stopes using Stability Graph Method
Stability Graph - Stope and Support Design Zones
The recommended limits for unsupported and supported stopes are combined,
along with the respective transition zones to obtain the design chart presented
below. This graph allows the engineer to determine, from a calculated value or
range of N', the maximum recommended stope size and shape for an unsupported
or supported case. A stope which plots well above or to the left of the uppermost
design curve is capable of remaining stable without support for a reasonable
service time. (Note that non-entry conditions are assumed here and that light
patterned rockbolt support and mesh may be required for personnel safety in other
areas). A stope which plots well into the lower-right quadrant is likely to suffer
major instability with or without support. The cablebolt design zone gives the
range in which cablebolts should be needed and effective. Clearly, the actual
effectiveness is reduced further right and down within this design zone. As HR is
increased or if N' deteriorates within this zone, the risk of failure is increased, and
standup-times are reduced requiring tighter cablebolt patterns and longer bolts.
Figure 2.17.10: Guidelines for cablebolt density
and length for regular patterns
(after Potvin, 1988)
Cable Support Recommendations - Potvin (1988)
Based on his original database, Potvin (1988) determined crude guidelines for
the design of patterned cablebolting. Specifically he proposed design charts for
cablebolt length and cablebolt density. Cablebolt length is the length of the
individual cablebolt (minimum length) and cablebolt density represents the
number of cablebolts per unit area of stope face.
For the design of cablebolt density, Potvin selected as the key empirical
parameter, (RQD/Jn)/HR. This represents a measure of relative block size with
respect to the excavation size. When this number was small it was expected that
an increased cablebolt density
would be necessary to ensure
stability. The resultant design
chart is shown at right. Note the
different zones shown here.
Based on this data set, Potvin
proposed that cablebolts were
ineffective when (RQD/Jn)/HR
was less than 0.6. In addition,
note that the practical minimum
cable density is 0.1 corresponding
to a square pattern of
approximately 3x3m. Three
cablebolt density design lines are
given which correspond to
different degrees of conservatism.
Non-entry stopes may require a
lower cablebolting density than a
main haulage drift for example.
The cablebolt length used in
each case study was plotted
against the hydraulic radius. This
follows logic based on classical
rules of thumb relating bolt length
and span. A representative line
based on current practice is shown
and corresponds approximately to:
Length = 1.5 x HR
up to a practical maximum of 15m
at a hydraulic radius, HR=10m.

Figure 2.17.11: Guidelines for cablebolt density to control local unravelling
Cablebolt Density (bolts/m of face) - Local Unravelling2
Potvin (1988) plotted cablebolt densities used in case histories against
(RQD/Jn)/HR based on the assumption that relative block size was in principle the
governing empirical parameter for stope face stability and support effectiveness.
Nickson (1992), however, applied statistical techniques in an investigation of
many possible parametric combinations. For the combined cablebolted stope
database of Potvin and Nickson, (RQD/Jn)/HR actually gave a very poor
correlation to cablebolt density based on current practice. This is illustrated by the
scatter in Figure 2.17.10. It is proposed here that the absolute block size
represented by RQD/Jn should control local block fallout from the face and
therefore should strongly influence ultimate stability of the stope. If cablebolts are
spaced too far apart, unravelling will occur between bolts, progressively leading
to more serious instability. The corresponding graph based on the Potvin/Nickson
database is shown in Figure 2.17.11. The design zone plotted provides a crude
recommended design range for cablebolt density in open stope applications. This
design zone should not be applied to permanent openings or in high traffic areas,
where safety is a critical issue, unless accompanied by primary support such as
rockbolts and screen.
Figure 2.17.12: Guidelines for cable spacing and density - overall stope face stability
Cablebolt Density or Spacing - Stope Face Support
Nickson (1992) showed that the best empirical correlation with respect to
cablebolt density was obtained by plotting density with respect to the parameter
N'/HR. The logic here is similar to Potvin's usage of (RQD/Jn)/HR, except that
N' contains additional information about stope inclination, stress related fracturing
(parameter A) and favourable or unfavourable joint orientations. Nickson derived
a relationship based on current practice without considering the degree of support
effectiveness. The design zones proposed above in Figure 2.17.12 do relate to this
degree of success. While the data scatter is great due to the trial-and-error nature
of present design practice, there appears to be a reasonable limit to cablebolt
effectiveness as delineated by the cluster of caved cases in the upper portion of this
plot. The non-conservative zone can be used as a guide for non-entry conditions
or where dilution is not critical. The conservative zone is applicable to stope backs
above drilling horizons and other areas where entry is permitted. Note the two
vertical scales used here. These illustrate the relationship between cablebolt
density and the cable spacing of an equivalent square pattern. Use Figures 2.17.11
and 2.17.12 together to determine the critical (maximum recommended) spacing.

Figure 2.17.13: Five design zones for cable support of open stopes
2.17.4 Semi-Empirical Cablebolt Design Approach
It is possible to combine the information gained from empirical methods with
mechanistic assumptions and logic to develop a more sound semi-empirical design
methodology. Figure 2.17.13 illustrates this approach. The No-Support Zone and
the Unsupportable Zone are derived as previously discussed from examination of
over 350 case histories. Within the cablebolt design zone (shaded area), however,
it is possible to assume some basic support functions and modify the design
accordingly. The Reinforcement zone implies that the rockmass is still partially
stable, requiring cables to merely hold together the constituent blocks to form a
self-supporting arch or beam. Cable spacings and lengths along the upper
boundary of this zone are derived directly from the analysis in Section 2.18.12
using a back calculated rockmass stiffness. In the Support zone, however, the
cables must bear the full load of the failed or loosened rockmass. Spacings and
lengths along the lower boundary of this zone are therefore derived from
conservative civil engineering experience (Section 2.16.5). The transition between
these two extremes is continuous across the shaded zone. Retention
recommendations based on ravelling failure (Figure 2.17.12) are superimposed on
the above results. The maximum spacing and minimum length required to
effectively carry out all of the support functions considered are then plotted in the
following charts. The zone marked Retain in Figure 2.17.13 is the zone in which
this function is critical with respect to spacing of cablebolt support. In the other
zones, reinforcement and support dictate the maximum allowable spacing.
Figure 2.17.14: Recommended spacing for single strand cables (regular pattern)
Maximum Design Spacing for Single Strand Cablebolts
Based on the assumptions illustrated in Figure 2.17.13, recommended cablebolt
spacings (for an equivalent square pattern) have been calculated for the range of
reinforcement-support across the shaded cablebolt support zone. Where the
maximum spacing so determined exceeds the recommended spacings obtained
from Figure 2.17.12, unravelling between and around cables is assumed to
dominate stability and Figure 2.17.12 therefore controls the design. The composite
result is the cablebolt spacing design chart shown below in Figure 2.17.14.
For a given value of N' and HR plotting within the shaded cablebolt design
zone, it is possible to determine the maximum (critical) spacing of single cables
in a square pattern to ensure stability. Note that minimum cablebolt density, Dc,
is related to maximum equivalent square spacing, Sc, as follows:
Cable Density, Dc (bolts/m ): Dc = Sc Sc = Cable Spacing (m)2 -2

Figure 2.17.15: Recommended spacings for double strand cablebolts
Maximum Design Spacing for Double Strand Cables
Single cablebolts (15.8mm strand) can be assumed to have 20 - 25 tonnes (200-
250 kN) of long term capacity provided that the bond strength and embedment
length are adequate. Double strand cables normally possess approximately twice
this capacity. Figure 2.17.15 below gives design ranges for double strand cablebolt
spacing. Again it is important to emphasize that full load transfer from the rock
to the cable is assumed. This implies good quality control and/or the use of
modified geometry cables (birdcage, bulbed strand, etc) and/or the use of plates
when practical. Note the expanded patterns as compared with single strand cables.
Also note that double cables make little difference in the lower-left retention zone.
Instability in this region is not related to steel capacity but only to interbolt
distance. Spacings can be increased as shown (dashed lines) when cables are used
in combination with a tight pattern of rebar or rockbolts or shotcrete. These
primary support elements serve to retain blocks and knit together a surface layer
which can be supported with an expanded pattern of cablebolts.
Figure 2.17.16: Recommended minimum lengths for grouted cablebolts
Minimum Design Length for Cablebolts (Single/Double Strand)
Support design at the outer limits of the Reinforcement and the Support zones
illustrated in Figure 2.17.13 are based on limiting conditions of arch/beam
reinforcement and deadload estimation respectively. Based on parametric analysis
using conservative parameters derived from N', these analyses yield the bounding
values for spacing discussed in the previous sections and for length as shown
below in Figure 2.17.16.
Recommended lengths for cement grouted cablebolts differ from resin grouted
or mechanical bolt recommendations in the literature. This is due to the necessary
addition of a reliable anchor length beyond the zone of supported rock. In the case
of beam analysis and deadload estimation, this corresponds in the figure below to
2m beyond the stabilized beam or failed zone respectively. Note that increasing
length does not always imply increased capacity (controlled by strand density).
These lengths are based primarily on cable coverage of the supported zone.

Figure 2.17.17: Regular Patterned Support
Cablebolt Spacing and Length of Regular, Uniform Arrays
All of the preceding discussion
concerning the modified stability graph
and recommendations for cablebolt
spacing and length apply to a regular or
patterned array of cables; a constant
distribution of bolts across the face area
of the stope and an arrangement behind
the face such that neighbouring cables
are within 40 degrees of being mutually
parallel. The example cablebolt patterns
in Figure 2.17.17 illustrate the ideal
application of these guidelines.
Cables should be spaced as close to
square as possible if designed using the
recommendations in this section. For
example, patterns such as 1.5m x 2m
(equivalent square = 1.7m x 1.7m) or
2m x 2.5m (= 2.2m x 2.2m) are
acceptable, whereas a pattern of 1m x
3m may not perform as well as the
equivalent square pattern (1.7m x
1.7m).
Cable spacing should also be uniform
(i.e. spacing should not vary more than
20% over the stope face). The density of
tight clusters of cables bounding larger
areas of unsupported stope face cannot
be averaged over the whole area and
equated to an average density or
equivalent spacing. The design of this
type of system is handled differently as
shown on the following page.
Cable lengths are specified for
cablebolts which are within 30-40
degrees of perpendicular to the stope
face. Normally the length refers to the
perpendicular distance between the face
and the end of the cables. Actual cable
length will depend on the cable angle.
Figure 2.17.18: Line anchor support system geometry
Line or Point Anchor Arrays - Sub-Span Design
In many cases in mining, access constraints do not allow the installation of a
regular uniform pattern of cablebolts in a back or hangingwall. In addition, mining
influences such as blasting, induced stress change or rock relaxation may limit the
effectiveness of a distributed cable pattern (Section 2.6). This is particularly the
case in foliated hangingwalls. Often it may be preferable, therefore, from both an
operational and an engineering viewpoint to install line anchors as shown below
at prescribed intervals. These anchors reinforce a local volume of rock, limiting
internal displacements and preventing dilation. This artificial rockmass block or
rib then acts as an effective abutment for adjacent spans (Fuller, 1983).
This support system should only be used in rockmasses dominated by a single
lamination parallel to the stope face or joint sets perpendicular to this face (few
oblique joints). Blast control is critical to avoid damage to the unsupported span
and displacement rate monitoring may be a useful design verification tool here.

Figure 2.17.19: Crude relationship relating overall (supported) span to unsupported
sub-span. Applicable to hangingwalls only (data from Nickson,1992).
As such these anchors must have a locally dense arrangement (1.5 m spacing
at collar) and 4-6 cablebolts in each ring. These cables should then be plated. This
is to ensure limited internal movement within this reinforced abutment. The
Modified Stability Graph can then be used directly to dimension the unsupported
sub-spans (a x b in Fig. 2.17.18). These sub-spans (unsupported spans) may be
strung together providing a huge operational benefit by allowing a much larger
stope to be opened without immediate backfilling. There is a limiting relationship,
however, between the unsupported sub-span and the overall supported span (or
hydraulic radius of total open stope face). Nickson (1992) compared 13 case
histories of line anchored hangingwalls and proposed the crude relationship
illustrated in Figure 2.17.19.
Note that the database is extremely limited and so caution must be exercised
when using this graph. Calibration to local conditions will be necessary.
The relationship above should not be applied to shallow dipping hangingwalls
or backs. This method is designed for non-entry stopes should not be applied to
stope faces in areas where regular human access is necessary without additional
primary support such as rockbolts and screen to control small block fallout.
2.17.5 Stability Graph - Examples
Consider the following examples of open stope scenarios. These four cases
have been deliberately chosen to result in the same hydraulic radii, HR and the
same values of Modified Stability Number, N'.
Table 2.17.2: Four example applications of the Stability Graph
CASE A
Hangingwall
CASE B
Back
CASE C
Hangingwall
CASE D
Back
Problem
Description
Depth 200 m 600 m 150 m 1000 m
Wall Stress 10 MPa 20 MPa 8 MPa 60 MPa
RQD 40 60 85 90
Joint Sets 2 2 + random 3 + random 2 + random
Joint Surface Smooth Planar;
Slightly Altered
Rough
Undulating;
Unaltered
Stained
Rough
Planar;
Slightly
Altered
Slickensided
Undulating;
Unaltered
Stained
Rock Type Foliated Schist Bedded
Limestone
Gneiss Massive
Sulphide
Rock Strength 80 MPa 115 MPa 160 MPa 180 MPa
Input
Parameters
Wall
Dimensions 20 m X 40 m 18 m X 55 m 25 m X 30 m 22 m X 34 m
RQD/Jn 40 / 4 = 10 60 / 6 = 10 85 / 12 = 7.1 90 / 6 = 15
Jr/Ja 0.5 3.0 0.75 1.5
A 0.78 0.52 1.0 0.21
B 0.3 0.3 0.3 1.0
C 8.0 2.0 6.0 2.0
Stability Graph
Coordinates
HR 6.7 6.8 6.8 6.7
N' 9.4 9.4 9.6 9.6
Status STABLE CAVED UNSTABLE CAVED

These examples are illustrated in Figure 2.17.20 on the following page. Note
the obvious differences in stope dimensions and in geometrical and geomechanical
environment. Yet the plotted results for the four cases are indistinguishable. This
illustrates both a strength and a weakness of the Stability Graph method. Like all
empirical methods it is a general design method which allows us to formulate
preliminary designs in the face of limitless variety and complexity. Other design
techniques are normally very problem specific and cannot be universally applied.
Once the preliminary design is established, however, the method does not provide
for the fine tuning which must occur to adapt the design to the specific problems
encountered in each case.
Case A, for example, is a thinly laminated rockmass with a second
discontinuous joint set at 90 degrees from the main lamination. Even though the
RQD is low due to the foliation, the stope wall is vertical and as such should be
inherently stable unless disturbed by poor blasting or excessive span development.
Cables are unlikely to improve stability within economic limits in this case. The
stress is low compared to the rock strength so gravity is likely to be a dominant
control. This case is suitable to Voussoir beam analysis (Section 2.18.12).
Case B is a competent blocky rockmass above a relatively wide sill span. The
main lamination would suggest beam analysis. The cross jointing, however, is
oblique to the back and is unlikely to allow complete arch development in the
horizontal back. Patterned roof bolting will be necessary in this case
Case C represents a strong gneiss with moderate structural density. Even though
the block size is small, the joint surfaces are very rough and tightly interlocked.
The stresses are low but the steep dip of the wall will maintain compression and
improve stability. Patterned cablebolting from a hangingwall drift should prove
effective in this case.
Case D appears to be the highest quality rockmass as indicated by the large
values for RQD/Jn and Jr/Ja. The stability graph analysis does not consider the
sheared contact which forms the hangingwall. It is likely that this contact will
shear due to stresses in over the back. These stresses are high and this slippage
may be unstoppable. The vertical jointing will form release planes resulting in a
large free full span wedge which must be supported. Cables must be designed to
withstand large displacements or they will snap as the wedge slips.
These examples show that while the Stability Graph method is an invaluable
tool for initial dimensioning and support design for open stopes, it is not the final
word. If the stope plots well into the stable zone or well into the caving zone, then
the respective result is fairly reliable. If the stope plots close to or within the
cablebolt design (support required) zone, then further mechanistic analysis should
be carried out to confirm the validity of critical assumptions and recommendations
of the stability graph method.
Figure 2.17.20: Four application examples for the Stability Graph Method. Note that very
different design problems can result in the same position on the graph.

Figure 2.17.21: Limitations for use of
Modified Stability Graph
2.17.6 Stability Graph Method - Limitations
There are certain assumptions
inherent in the application of the
Stability Graph Method. Observe the
following limitations when using the
method for stope and support design.
Inadequate Fill
The estimation of stable hydraulic
radii determined from the graph or used
as input for stability evaluation assumes
that the span being considered is fully
bounded. This assumption is valid for
unfilled stopes which are surrounded by
fill (as in alternate block sequencing, for
example). The surrounding fill must be
tight to the walls and back of the stopes
in order to be considered supporting
elements. If this is not the case as
shown in Fig. 2.17.21a), the true
effective span for analysis may be much
larger than the nominal stope panel.
The same is true if the fill is highly
compressible. In such a case, the
Stability Graph Method is not applicable
to the design of the unmined panel 7 in
Fig 2.17.21b), for example.
Corners-Designed and Accidental
As shown in Fig 2.17.21c), corners
or bulges can be created in stope walls
though poor design (chasing grade) or
though the upward caving of mined
stopes below. In either case, the stability
graph cannot be used to evaluate the
stability of either the span above the
corner nor the overall span. The corner
so created, will dominate the stability of
the entire stope and will likely cause
major stability problems. Such corners,
either deliberate or accidental should be
avoided.
Figure 2.17.22: Limitations of the
Stability Graph Method
Intersections
While the Stability Graph Method
was calibrated for open stopes, it can be
used for large mining tunnels and sills
provided that a conservative approach is
taken for safety reasons. The method
should not be used, however, to design
intersections. The assumption of a
bounded span is not valid here.
Intersections are normally less stable
than the associated tunnels. In addition,
it is not possible to calculate an
equivalent HR for an intersection.
Discrete Wedges
The stability graph design approach
is applicable to moderately structured
rockmasses with distributed or
ubiquitous structure. Discrete structural
features such as large wedges which
may form in sill backs must be
considered separately.
Delamination Zones
Large stable spans may be predicted
in cases with structurally sound wall
rock. If this wall rock is bounded by a
weak layer close (within 20% of the
span) to and parallel to the wall as
shown at left, the stability of the
resultant beam will be reduced. Beam
analysis methods may be more
appropriate for design.
Discrete Shear Structures
Large scale structure (length stope
dimensions) will control stope stability
under stress and gravity. Discontinuum
analysis methods must be used for
design. The Stability Graph results will
not accurately predict stability.

Figure 2.17.23: Example of local site calibration. New design line (dashed) can be used
for future stope design (after Bawden, 1993).
2.17.7 Stability Graph-Calibration to Local Conditions
The initial database of approximately 350 case histories is impressive in size
and scope. It is, however, incapable of accurately predicting stability in every
possible situation. The database, for example, reflects Canadian practice. This
immediately implies a bias towards Canadian conditions. In short, the method
provides an excellent starting point for design but it must be calibrated on-site in
every new mining environment. This involves maintaining an up-to-date database
of stope dimensions, rockmass parameters, and stability status.
Bawden (1993) uses a data set from Greer (1989) to demonstrate this concept
as illustrated in Figure 2.17.23 below (note the truncated axes for more detail). In
this case, due to unique conditions at the mine site, significant caving and
instability was observed in stopes which the method predicted to be stable.
If such a local database is maintained, then the Stability Graph can be calibrated
for local conditions. The dashed design line proposed by Bawden for the above
data bounds the caved and unstable stopes. This line should now be used as the no-
support limit for future mine design at this site.
2.17.8 Parametric Analysis
The quality of a rockmass is never well defined. For overall mine design and
budgeting purposes at a preliminary stage it may be adequate to design based on
average rockmass conditions. Assuming worst case parameters may prove
impractical from an economic perspective while designing based on the best
possible conditions would clearly be imprudent.
In order to understand the consequences of this variability at a given site, it is
useful to employ a bounded analysis for the Stability Graph method by tabulating
reasonable ranges for the input parameters (limit ranges to one tabulated category
or one standard deviation for each parameter and only use variability as required
or impractically large solution ranges will result) and then calculating an expected
range for N'. Consider the following example input for the hangingwall of a mine
employing a modified AVOCA mining method:.
Table 2.17.3: Data sheet for parametric design example
DEPTH 500m
STOPE HEIGHT 20m
NOMINAL PANEL
WIDTH
30m
WALL
STRESS
20 - 30 MPa
HW DIP 65 degrees
HW ROCK gneiss
UCS 120 - 180 MPa
JOINTS 2 + random rough planar stained -
slightly altered
PARAMETER LOWER BOUND EXPECTED UPPER BOUND
RQD 70 75 80
Jn 6 6 6
Jr 1.5 1.5 1.5
Ja 2 2 1
A 0.3 0.4 0.5
B 0.2 0.3 0.4
C 5.5 5.5 5.5
N' 2.9 6.2 22

20 30
40 60
6
×
+
=
20 45
40 90
7
×
+
=
Figure 2.17.24: AVOCA Stope example - Fill Lag can be up to 1/2 stope width
Figure 2.17.25: Example design range and recommendations
Due to operational delays and scheduling problems, it is expected that the
backfill front as illustrated in Figure 2.17.24 can lag behind the design position
(relative to the blasting face) by as much as 1/2 of the nominal panel width. It
must be considered then that the width can vary from 30m to 45m. The hydraulic
radius, HR, must therefore be assumed to vary in the range:
From: To:
The combined range of N' and of HR can be plotted on the Stability Graph as
shown in Figure 2.17.25. Support is clearly required in this case. When worst case
conditions occur, significant stability problems may result if support is inadequate.
The decision to enhance support beyond average requirements must be based on
risk to personnel and equipment and on the costs, losses and delays associated with
unexpected dilution. This method can be expanded to involve probabilistic
methods similar to those outlined in Hoek et al. (1995) and Harr (1987).
Figure 2.17.26: Example input distributions resulting from in situ variability
Figure 2.17.27: Probabilities of (b) instability and caving based on (a) Monte Carlo
analysis (after Diederichs and Kaiser, 1996)
2.17.9 Probabilistic Analysis
Another approach to incorporating input variability into the analysis is through
the use of a probabilistic approach. Input parameters can be assigned distributions
as shown in Figure 2.17.26. If a single recorded value for any parameter is
sampled at random from the database, these histogram distributions show the
relative likelihood of the sample equating to a particular value or range.
Distributions can be obtained from real field data using statistical techniques
(Harr, 1987; Pine, 1992; Rosenbleuth, 1981) or commercial simulation software
(Hoek et al., 1995; Carter, 1992; Diederichs and Kaiser, 1996), and can be used
in a Monte Carlo style analysis. In this analysis, a large number of calculations for
N' are generated from different combinations of values for the above parameters.
The frequencies with which each parameter falls within different ranges for use
in the calculation are reflected in the distributions in Figure 2.17.26. Several
hundred such calculations result in the distribution for N' shown in Figure 2.17.27
a). If this distribution is superimposed on the instability limits at a given HR, the
probability of instability or of caving (Figure 2.17.27.b) is equal to relative area
of the distribution which falls below the respective limit.

Figure 2.17.28: Site-specific average expected dilution (data from Pakalnis et al., 1995)
2.17.10 Dilution and the Stability Graph
The instability and caving limits in the Modified Stability Graph are based
loosely on the apparent area of instability across the stope face. If the volume of
failure is considered and divided by the volume of the ore in the stope, a value for
dilution is obtained (Section 1.2). For a simple rectangular geometry, and if the
stope thickness does not change, it is possible to plot contours of expected average
dilution on the Stability Graph (Figure 2.17.28). Note that these contours are
likely to be site-specific and depend on the stope thickness (5m in the example
below). Based on local site experience, a dilution vs HR relationship for any rock
quality N' can be obtained and used in economic analyses to optimize stope
dimensions (Elbrond, 1994; Planeta et al., 1990; Diederichs and Kaiser, 1996).
2.18 A Mechanistic Toolbox:
Customizing the Design
While empirical design methods typically produce general preliminary
recommendations to cover a wide variety of rockmass behaviour (within a given
rock quality range), a mechanistic approach considers specific failure mechanisms
and adjusts design accordingly. Hoek and Brown (1980), Hoek et al. (1995),
Brady and Brown (1993), and others give detailed treatment to many of these
mechanisms and to the appropriate support strategies. A modest selection is
covered here.
Rockmass data collection
Rockmass characterization and classification
Identification of potential failure modes
Structurally controlled, gravity
driven blocks and wedges
Stress induced, gravity assisted
failures
Evaluation of kinematically
possible failure modes
Determination of in situ stress
field in surrounding rock
Assignment of shear strength to
potential failure surfaces
Assignment of rockmass
properties
Calculation of factor of safety or
risk of potential failures
Analysis of size of overstress
zones around excavations
Determination of support
requirements
Non-linear support-interaction
analysis to design support
Influence of blasting, dynamic disturbance, and time dependency
Design of support, taking into account excavation sequences, availability of materials
and cost effectiveness of the design
Support installation
Monitoring
Figure 2.18.1: Mechanistic Design (Italics). After Hoek et al. (1995)

2.18.1 Stress Induced Boundary Crushing
Hoek et al. (1995) describe a methodology for analysis of excavation induced
stresses for the purpose of support design. Two-dimensional or three-dimensional
elastic analysis may be used to evaluate the induced stresses around complex
excavation shapes. These stresses can be compared to an appropriate strength
criteria to determine the extent of rockmass failure (Sections 2.13 and 2.15).
In hard and brittle rockmasses, the maximum compressive stress can be
contoured around an excavation boundary (where the minor or least compressive
stress approaches zero) and compared with the uniaxial compressive strength of
the rock. Where the calculated compressive stress exceeds one-half (Section
2.13.3) of the strength determined from testing of laboratory samples, it can be
assumed that, in the long term, the rockmass will become significantly damaged
and may require support. This method of analysis is most appropriate in highly
stressed, hard, brittle rockmasses with moderate to low initial fracturing or
structure. Failed rock in these environments is unlikely to possess much residual
strength and will require full support after the creation of the failed zone. For
regions within one excavation radius, compare the induced stress difference (. -1
. ) to 0.5 times the laboratory UCS. Areas where the stress difference exceeds this3
value may be prone to damage and eventual weakening and rupture.
In softer and more fractured rockmasses, the use of a confinement dependent
criteria such as Hoek-Brown (Hoek et al., 1995) is warranted. Plastic analysis may
be used to assess the potential for progressive failure and to investigate the role of
stress redistribution and self-stabilization. Once a zone of potential failure has
been established from such models, it may prudent to simply design cablebolt
support to sustain the deadload of this failed zone (Factor of Safety or Strength
Factor 1). Often this will prove economically impractical and may not be
necessary, since failed rock often retains limited ability to support itself. If plastic
analysis has been employed, the support must extend into the zone of confinement
or the zone above the back (e.g. in a roof support problem) where the minor (least
compressive) stress begins to increase. This indicates the development of self-
stability. The location of this boundary will, however, be highly dependent on the
strength and deformation parameters used in the analysis.
It should be noted in any case that cablebolts are unlikely to arrest the onset of
rock failure under high stress and may do little to alter the progression of such
failure into the rockmass. The objective here is to hold the failed material in place
so that the broken rock itself can generate the necessary confinement to reduce the
extent of progressive damage and instability. In highly plastic (deformable)
rockmasses under high stress, it is also unlikely that cables will be effective in
arresting the progression of failure. In addition, in these environments, the induced
displacements may be too great for the system to handle and cable strand rupture
may be inevitable in pre-installed systems.
2.18.2 Stress Shadowing and Relaxation
Low stresses can pose as much risk as high stresses in a fractured or jointed
rockmass. When confinement is present, joint surfaces remain mated and even
limited surface roughness provides adequate dilational and frictional resistance to
slip. When joints are subjected to shearing in a confined environment, the
asperities on the surface interact and interfere with each other as slip progresses.
This results in a tendency for the joint to open which in turn generates increased
pressure on the asperities, increasing the frictional resistance to further slip. The
only avenue for further slip becomes shear through of the asperities requiring large
stresses. In this way, joint slip under confining stress may be self-limiting.
When confinement is removed as in a late stage stope boundary after significant
stress and blast damage has occurred and after neighbouring excavations have
provided a stress shadow, the fractures are free to dilate and slip resulting in a
destabilized rockmass. While even modest stresses across a back or sidewall can
serve to clamp the rockmass blocks in place, gravity will dominate once these
stresses decrease.
In an elastic model, zones of tension or zones of near-zero stress in, for
example, a stope wall indicate potential problem areas (refer to Figure 2.13.10).
This applies to highly fractured rockmasses, blocky rockmasses and also to areas
where discrete, large scale structure is present. Large wedges and blocks can be
liberated in otherwise stable areas by reductions in field stress.
At particular risk are zones of rock which have previously undergone overstress
failure (confined) and which are, at a later stage in mine life, undergoing
relaxation. The previously self-stable rockmass may tend to unravel and cave in
this situation. Cablebolt support in these areas can be designed to support the
deadload of such destressed zones. Roof bolts in cut and fill stopes, for example,
must be designed in this way. Where the ore is weak, ensure that the cables
penetrate into the hangingwall (Cassidy, 1980). If this is not possible, then
approaches such as beam building (Voussoir) can be employed to optimize design.
In any case, it is usually prudent to design cable lengths to penetrate into a zone
of confinement.
Zones of relaxation pose an additional hazard for cablebolting. As discussed in
Section 2.6, stress decreases across a cable array can seriously impair the bond
strength of plain strand cablebolts. Rockmass stiffness is also dependent on
confinement in fractured rockmasses, and decreases with relaxation. This has a
compounded detrimental effect on cable capacity - just when bond strength is
needed the most.
It is for this reason that plating and the use of modified strand cablebolts are
recommended in fractured-destressed rock.

Figure 2.18.2: Effect of displacement on rockmass integrity
2.18.3 Limiting Displacement - Reinforcement
Undisturbed, fractured rockmasses in underground environments are generally
inherently stable, provided that excavation can proceed with little or no additional
disturbance and if internal displacements (block or particle shifting) can be
minimized. Rockmass quality, stiffness and strength generally degrade with
displacement.
Consider the example illustrated in Figure 2.18.2. As the hangingwall displaces
(at midspan) the constituent blocks shift with respect to each other and lose the
essential interlocking required for stability. Rockmass joint surfaces typically have
a characteristic roughness scale (the basis for Jr/Ja in the Q system for example)
or asperity height. This determines the degree of interblock slip which can be
tolerated before significant interlock is lost. Rockmasses will rapidly disintegrate
at this time as shown in Figure 2.18.2. This behaviour suggests a limiting
displacement which can form the basis for cablebolt pattern and element type
selection. In general, rockmasses with small block sizes or low roughness fracture
surfaces will have smaller critical displacements than large block size, rough
jointed rockmasses.
When the limiting displacement is small, such as in the case of a highly
schistose hangingwall, a very stiff system such as multiple bulbed strand on tight
spacings may be required.
This critical displacement can be determined by systematic monitoring from
stope to stope (Chapter 4). Design can be optimized in this way.
Figure 2.18.3: Depth of failure for
rockmasses with varying ductility
Figure 2.18.4: Relative surface
displacement vs rockmass ductility
Displacements in Hard Rock
The yield or damage strength and
rupture strength of hard rock can be
several orders of magnitude higher than
the shear strength contribution of a
distributed pattern of cablebolts. It is
unlikely therefore that cablebolts can
prevent the onset of fracturing and
damage in hard rockmasses. In practice,
however, cablebolted excavations in
overstressed rockmasses normally
perform better than if left unsupported. It
is well established in rock mechanics, that
ductile rockmasses (Figure 2.13.11.b) are
more self-supporting after yield than
brittle rockmasses of equivalent peak
strength. The role of cablebolts in a hard
and brittle rockmass must therefore be to
restrain the dilation of existing
discontinuities and of stress induced
fractures so that the post yield strength of
the rockmass is maximized and ductility
is achieved. Open or dilating joints or
fractures can drastically reduce the
ultimate strength of a rockmass.
Analyses by Kaiser (1980), show the extent of the failure zone (Figure 2.18.3)
and the surface displacements (Figure 2.18.4) around a circular opening for
different degrees of rockmass ductility (s = ultimate strength/peak strength; s=0
for perfectly brittle and s=1 for perfectly ductile) for an example rockmass.
Limiting displacement through support creates a more ductile rockmass which
carries load after yield, and maintains confinement away from the opening,
limiting the extent of failure. This, in turn, further limits the total displacement.
Displacements in Soft Rock
In a soft and weak rockmass, cables perform in a similar fashion as described
above. A yielding rock such as phyllite or schist can form either discrete or
distributed slip surfaces throughout the yielding zone. These opposing surfaces
must dilate (open or separate slightly) in order to shear past each other.
Suppressing this dilation increases the shear strength. Single discrete and persistent
surfaces require less total dilation to slip and therefore shear more readily than a
combination of many small distributed surfaces. Cablebolts act to restrain discrete
weakness planes in the soft material. Yielding can then only occur evenly and
continuously throughout the rockmass. Even a weak and yielding continuum is
stronger than a discretely discontinuous rock of similar material properties.

Figure 2.18.5: Stress induced slip along smooth discontinuities
2.18.4 Stress Induced Joint Slip
Where slickensided planar joints intersect an excavation boundary at a low
angle of between 5 and 45 degrees, there is potential for stress driven displacement
or slip across these surfaces. The driving force behind these displacements will
normally be beyond the capacities of conventional support in hard rock at depth.
These movements are normally shortlived, however, as the stresses are relieved
by the slip and transferred out into the rockmass. The goal of pre-support, in this
case is to go along for the ride during initial slip, maintaining load capacity to
support the relaxed blocks created by this deformation. In a highly stressed
environment with several unfavourable shears converging at an excavation
boundary, a required displacement capacity of 100mm would not be extraordinary.
Once induced stresses around the excavation are obtained using elastic analysis,
the procedure summarized in Figure 2.13.7 in Section 2.13.2 can be used to
determine the tendency for inclined discontinuities to slip or to be clamped by the
induced stress field.
This analysis can be used in conjunction with simple
gravity wedge analyses (Sections 2.18.8 and 2.18.9).
Wedge analyses can often reveal enormous and steep
sided wedges in the back of an excavation. This can be
cause for alarm. It can be shown, however, that even
with a modest degree of lateral confinement ( 1 Mpa),
a rock wedge with apex angle, , less than the effective
friction angle (Section 2.13.3) will be stable (i.e.
N+i). This explains why, at depth and in hard rock with
rough joint surfaces (i.e. N+i 45-), failures involving
wedges with heights, H, greater than 1/2 of the
excavation span, S, are rare (i.e. tan(45)=1; XH =0.5S).max
Figure 2.18.6: Dynamic Loading
Figure 2.18.7: Composite cablebolt - rebar
support (Plain strand, plated cablebolts can
also be used in less severe conditions.)
Figure 2.18.8: Cable-Swellex bolt
2.18.5 Dynamic Loading
Support displacement capacity is critical
in areas of seismically induced or blast
generated dynamic loading (Ortlepp, 1983;
Kaiser et al., 1995; Kaiser and McCreath,
1992). Cablebolts can be effective where
fractured ground has relaxed and become
vulnerable to vibration. Support loads
generated over larger displacements absorb
more of the kinetic energy of such a ground
mass. Stiff supports will rupture before the
mass can be decelerated and stabilized.
It may be advisable to combine long,
partially debonded cables, with modified
geometry or plated end lengths, with
closely spaced primary support such as
grouted rebar (Kaiser et al., 1995). Under
dynamic loading the rebar will maintain
a reinforced skin at the excavation
surface. The large displacement of this
skin can be accommodated by the cables,
maintaining ultimate holding capacity
after the disturbance has passed. Without
the rebar, the surface skin may loosen and
disintegrate, rendering the cablebolts
ineffective.
Tannant and Kaiser (1995) describe an
innovative support element for support of
blocky ground under dynamic loading. A
cable is anchored in a borehole using a
Swellex friction bolt (Stillborg, 1986). A
length of cable extends beyond the
downhole extent of the Swellex and has an
optional button or swage (Section 2.9.3) at
the end. Even if the Swellex breaks in
several places under loading, the cable
remains frictionally anchored within the
Swellex segments. The load-displacement
characteristics are ideal for dynamic
conditions; a consistent yield load and a
large displacement capacity before rupture.

Figure 2.18.10: Bolt support for a sliding wedge (after Choquet and Hadjigeorgiou,1993)
Figure 2.18.9: Surface retention
2.18.6 Surface Unravelling
The spacings for cablebolt elements
as calculated in designs based on
deadload and support pressure are
greater than those calculated for rock
bolts, due to the higher capacity of the
cables. In addition, cablebolts are not
tensioned (although proper plating
provides up to 5 tonnes of loading on
the surface of the rock - comparable to
a rockbolt). The increased spacing
between pattern cables and the lack of
active loading (unplated cables)
provides a greater opportunity for
blocks of rock between the cablebolts to
fall freely from the surface. In highly
fractured rock, supplementary support
and face restraint systems will have to
be used in conjunction with cablebolts.
2.18.7 Sliding Wedge
Figure 2.18.11: Bolt spacing and total load for a prismatic gravity wedge
Figure 2.18.12: Critical embedment length adjustment
2.18.8 Two-Dimensional Wedge
In many inclined tabular ore bodies, it is common to find two dominant joint
sets; one parallel to the hangingwall (or footwall) and one cutting obliquely across
the back. Vertical release planes normal to the ore zone complete the definition of
two-dimensional prismatic wedges. These wedges can and often do form across
the full span of the stope and must be supported. Consider the bolt spacing and
total capacity calculations in Figure 2.18.11.
When unplated cables are to be used for this application, it is important to be
aware of stress change effects inside the wedge which may lead to bond capacity
loss (Section 2.6). In addition, all of the cables within a regular pattern will not
have the same load response curve when supporting a gravity wedge. Cables on
the outer edges may not have embedment lengths greater than their critical
embedment length (length required to break steel during pullout). These cables
will have a much softer response and will not take on load at the same rate as the
other cables towards the centre of the wedge. As a result, the centre cables may
become overloaded and fail in tension even though the total tensile capacity of all
of the cables may be more than the weight of the wedge. It may be necessary to
move the outer cables closer towards the centre (use rockbolts at the wedge
perimeter), to plate the cables or to use modified strand.

Figure 2.18.13: Visualization of excavation and three-dimensional wedge with cablebolt
pattern support
2.18.9 Three-Dimensional Wedge
The analysis of three-dimensional wedges is based on limit equilibrium analysis
methods similar to the two-dimensional wedge analysis discussed in the previous
section. The definition of arbitrary wedges, based on joint orientation
information, and the calculation of weights and safety factors is more complex
than in two-dimensions and therefore routine treatment by hand, while possible
(Hoek and Brown, 1980), is not normally practical. Computer software is available
to perform the necessary stability calculations.
Three-dimensional wedges are easily visualized using the program UNWEDGE
(Hoek et al., 1995). The program allows the calculation of the weight of the
wedge, the input of support elements and calculation of the factor of safety against
failure (support capacity / gravity demand) using limit equilibrium techniques.
The output of the program is similar to that shown in Figure 2.18.13.
The same critical embedment length criteria for non-plated cablebolt support
of two-dimensional wedges also apply to the design of support for three-
dimensional wedges. A denser pattern in the centre of the wedge may be required.
σ
π
b =
2
2
12
E
S
T( )
Figure 2.18.14: Buckling Analysis
2.18.10 Stress Induced Buckling: Euler approach
In highly anisotropic (foliated) ground at high stress the foliation produces thin
slabs of rock which may be parallel to the excavation wall. These slabs can
potentially fail at a stress level much lower than the compressive strength of the
rock due to the instability phenomenon of buckling, in much the same way as a
thin sheet of strong steel will bow and collapse under a minimal load parallel to
the sheet. The calculation of a critical stress required for buckling is based on slab
geometry and rock stiffness and with the following assumptions:
+ The out-of-plane dimension of the slab is the largest dimension.
+ The thickness of the slab can be defined as the minimum spacing of potential
foliation parting of laminar joints.
+ The foliation slabs are sufficiently intact to warrant the use of the intact
modulus and compressive strength for stability calculations.
+ There is no inclined cross structure which could lead to alternate failure modes.
E = Intact rock stiffness (parallel to foliation)
S = In-plane slab span (long dimension)
T = In-plane slab thickness (short dimension)
The role of stiff or tensioned cables in this situation is to reduce the active spans
(i.e. S' = S / N; N is the number of evenly spaced bolts across the span).

Figure 2.18.15: Drift Support
Figure 2.16.16: Vertical deflection
of tunnel and intersection roof
2.18.11 Drift and Intersection Support
Cablebolts can be used to enhance primary
support systems such as rockbolts or grouted rebar
and screen, or shotcrete in drifts and tunnels.
While primary support systems hold the surface of
the excavation together, cablebolts perform the
function of tying this reinforced skin back into
sound rock. Cablebolts also provide added
security under seismic conditions.
When cablebolts are deemed necessary for drift stability, ensure that their
length is at least 3 metres beyond the maximum extent of yielding rock or beyond
the apex of the maximum expected rock wedge. The design load capacity of the
cablebolt system depends on the expected support function. If the cablebolts are
expected to carry a full deadload of yielded rock, then the load capacity should be
equivalent to a slab thickness of 30% to 50% of the span. This can be reduced to
10% to 20% of the span if the primary reinforcement can be expected to
successfully create a beam (Section 2.18.12). The empirical tunnel support
guidelines in Section 2.16 should also be of assistance. For excavations at depth,
follow the logic described in Section 2.18.1 and in Section 2.13.3. Hoek and
Brown (1980) present charts of induced stress, around various excavation shapes
and in various stress fields, for comparison to rock strength. Detournay (1988)
calculates depth of yielding around a circular opening in non-uniform stress fields.
Wedge analysis in structured ground is described in Sections 2.18.8 and 2.18.9.
Stability and support design in laminated ground is described in the next section.
Intersections pose a special design problem.
A major function of support is to transfer
rock loads from the centre of the span to the
abutments. At an intersection, the definition of
the abutments is not immediately apparent and
so additional logic is needed. Figure 2.18.16
illustrates typical vertical displacements along
the centreline of a tunnel approaching an
intersection. The displacement in the
intersection is equivalent to that in a tunnel
1.75 times wider than the actual tunnel span.
Since displacement ultimately affects stability,
the following rule of thumb can be adopted.
Wedge stability analysis can be assessed directly using geometric techniques
(Hoek and Brown, 1980) or using software as in Section 2.18.9. For other failure
modes, analyze an intersection roof as an equivalent tunnel 1.75 to 2 times the
maximum adjacent tunnel span or as a square with a width 2.5 to 3 times the span.
Alternatively, for complex shapes, extend the entire perimeter of the intersection
radially outwards by 1 tunnel span for the purposes of stability analyses.
2.18.12 Gravity Bending/Buckling:
No-Tension Slab - Voussoir Approach
Open stope mining often involves excavation of large span openings parallel
to laminated or bedded structure. When the discontinuity forming the lamination
is the sole structural feature to be considered and the excavation technique is
selected to minimize damage to the rock, standard beam analysis can be used to
evaluate the stability of the roof or hangingwall. It is more often the case,
however, that structures crosscut the main laminations reducing or eliminating the
ability of the rock to carry tensile loads parallel to the lamination making standard
elastic beam analysis inapplicable. In this case an alternative technique must be
employed.
The problem is similar to the design and
construction of mortarless masonry arches. The
solution technique, which has become known as
Voussoir analysis was first applied in rock
mechanics by Evans in 1941. It has been
modified over the years (Beer Meek, 1982;
Brady Brown, 1985), correcting some earlier
assumptions and improving the solution
technique. The solution presented here represents
a further development by the authors in order to
correctly incorporate arch deflections and to
incorporate more acceptable design confidence
limits into the solution.
Assumptions:
+ The out-of-plane depth of the beam is very large compared to the in-plane span.
Only a unit depth is considered with all deformations occurring in plane.
+ Cross-cutting structure is angled from the wall normal at significantly less than
the minimum angle of friction assumed for the jointed surfaces.
+ The beam is not capable of sustaining tension.
+ As the beam deflects, a parabolic compression arch develops in the beam.
+ Deflection of the beam occurs before slippage at the abutments. Stability
against slippage is determined after the compression arch develops.
+ Initial lateral stress resulting from in situ stress and excavation geometry is not
considered in this analysis. The beam is assumed to be initially stress free.
+ The abutments are stiff - they do not deform under the arching stress. For large
span to lamination thickness ratios, the deformation of the abutments can
normally be assumed to be negligible compared to the shortening of the roof
beam.

Figure 2.18.17: Problem geometry for Voussoir stability analysis
The problem is statically indeterminate. This means that there is no explicit
solution and that the iteratively obtained solution is approximate. A factor of
safety of 1.5 to 3 is advisable. In addition, the solution is highly sensitive to
rockmass modulus. The lowest expected value should be used. For a horizonal
beam, the problem geometry is as shown below.
The following input parameters must be specified:
E = Rockmass stiffness (parallel to excavation surface), (MPa)
UCS = Uniaxial compressive strength of rock, . , (MPa)c
S.G. = Specific Gravity, (dimensionless) or
= Specific weight of rock
T = Thickness of continuous laminations parallel to surface, (m)
S = Span of excavation surface being analyzed, (m)
In the case of a long excavation, S, is the short dimension.
= Inclination or dip of excavation surface, (degrees from horizontal)
Two failure modes are analyzed:
a) Crushing at the top and bottom of
the beam resulting in beam failure
when the compressive strength of
the rockmass is exceeded.
b) Snap-thru at the middle of the
beam resulting in immediate
collapse. Controlled mainly by
geometry.
Both failure modes are dependent
on inclination and density and are
most sensitive to rockmass modulus.
Figure 2.18.18: Calculation flow chart for the iterative Voussoir solution. Auxiliary
variables include: z=arch thrust moment arm between centre and
abutments); Fm, Fav=maximum and average arch stress; L=arch
shortening; N=ratio of arch thickness to beam thickness (0 to 1.0).
Note that the Buckling Limit = proportion of unsolvable cases for N.
Calculation Procedure

Figure 2.18.19: Limiting beam deflection for buckling and crushing failure modes
Deflection and Stability
Previously documented presentations of this solution have used an absolute
snap-thru limit which is defined as the limit of stable deflection according to the
mathematical formulation. This limit (Buckling Limit = 1) is extremely sensitive
to lamination thickness, a difficult parameter to reliably estimate and one which
may change as deflection and layer separation occurs. As a result, large safety
factors have been recommended (Beer and Meek, 1982; Brady and Brown, 1985).
The stability charts which follow utilize a design limit for snap-thru which is
based on a sensitivity or design confidence limit equivalent to a Buckling Limit of
0.35 in Figure 2.18.18. Beyond this limit (i.e. 0.35 to 1.0), small differences in
thickness have an unacceptably large influence on stability. As a result of this
adjustment, the charts which follow can be used with greater confidence in design.
Figure 2.18.19 above and Figure 2.18.20 also illustrate an interesting
component of the analysis which becomes useful for excavation monitoring and
design verification. Notice that for any span, inclination or rock modulus:
+ The design snap-thru limit is reached when midspan displacements reach 10%
of the lamination thickness. Beyond this deflection as in the case of example
A in Figure 2.18.19, stability is unlikely.
This critical displacement (deflection at failure) may be further reduced by low
compressive strength of the rock as crushing failure becomes dominant. Actual
midspan displacement at equilibrium for a stable excavation surface is dependent
on all of the input parameters (see example point B in Figure 2.18.19).
Figure 2.18.20: Limiting deflections for a variety of beam configurations
Deflection and Stability (cont)

Figure 2.18.21: Critical (maximum stable) span for laminated horizontal backs
Span vs Thickness: Horizontal surfaces
Considering separately the two failure modes of snap-thru and crushing, design
charts can be obtained as shown below for a horizontal excavation surface. A
stable span plots below the design curve for the appropriate rockmass modulus,
E (snap thru), and for the appropriate compressive strength, UCS (crushing). InRM
these charts, specific gravity is constant (S.G.=3.0) and b is the tunnel length.
Figure 2.18.22: Critical spans for laminated hangingwalls inclined at 65 degrees
Span vs Thickness: Inclined surfaces
Voussoir analysis can be applied to inclined surfaces as well. Certain
simplifying assumptions must be made which do not consider the distribution of
pressures due to self-weight acting parallel to the beam. Nevertheless, a reasonable
solution may be obtained and applied with the appropriate factor of safety ( 2).
The following charts are for laminated walls inclined at 65 degrees.

Figure 2.18.23: Effective specific gravity for generalized Voussoir analysis
Span vs Thickness: General solution
The method can be generalized for any inclination and for any specific gravity.
First calculate an Effective Specific Gravity, S.G. , based on the actual specificeff
gravity of the rock and on the inclination as shown below in Figure 2.18.23.
The next step is to obtain the normalized modulus E' by dividing the actual
modulus by the effective specific gravity. Next, the normalized compressive
strength UCS' is obtained by dividing the real UCS by the effective specific
gravity. Finally the maximum stable span for a beam can be found from the
assumed thickness using Figure 2.18.24. Note that this chart and those on the
preceding pages are applicable to a long stope wall with one dimension
significantly longer than the other. The span used in the analysis is the short span.
The results will be conservative.
A solution can also be obtained for a square stope surface (Brady and Brown,
1985). In this case, all four abutments contribute to the confinement of the beam
or plate. As a result this analysis will give less conservative results (e.g. larger safe
spans). The results for the general analysis are also given in Figure 2.18.24. Note
that both crushing and snap-thru failure modes are combined on each plot.
Most excavation spans will be rectangular. These two charts, therefore, serve to
bound the actual solution. Use both to obtain an upper and lower bound design.
Note that these charts are based on a Buckling Limit of 0.35 (Figure 2.18.18).
Critical spans based on a Buckling limit of 1.0 as in Beer and Meek (1982) and in
Brady and Brown (1985) will be up to 20% larger.
Figure 2.18.24: General solutions for Beam (infinite depth) and Square plate
Critical Span vs Thickness - General Voussoir Solution

Figure 2.18.25: Cable spacing and length guidelines using Voussoir approach
Support Rationale - Voussoir beam
If the assumptions inherent in this analysis can be validated, it is possible to
develop a support pattern to create a laminated beam or plate which will then
prevent further destabilization of the wall. The role of cables here is two fold.
Firstly, cables near the abutments act to reinforce the joint surfaces, increasing
resistance to internal shear which could lead to delamination and destabilization
(smaller thicknesses have smaller critical spans). Topsill and bottomsill cable
arrays (not shown in Figure 2.18.25) perform this role (Bywater and Fuller, 1983).
Secondly, cablebolts installed normal to the laminations and covering the span
area should be designed as stiff reinforcement within the zone of rock equivalent
in thickness to a self-supporting beam as calculated by this analysis. This is to
prevent delamination though the central portion of the beam (Roko and Daemen,
1983; Stimpson, 1983; Snyder, 1983). Beyond this limit, an optimum cable array
would have a more ductile response to allow the beam to deflect a small amount
to generate the required compression for stability. Beyond this should be a suitable
anchor length. Spacing as shown in Figure 2.18.25 is based on the deadload of the
beam. If the cables can hold the weight of this beam, then stability should be
assured. This result is usually much more efficient than a pure deadload estimate
on a relaxing hangingwall (no beam formation).
Figure 2.18.26: Mandolin bolting approach used
in Australia (after Cutjar et al., 1985)
Figure 2.18.27: Mandolin bolting in combination
with conventional cablebolt patterns (after
Lappalainen and Antikainen, 1987)
2.18.13 Other Applications
Mandolin Bolting
Cablebolts are ideally suited for axial loading. For this reason cablebolt patterns
are normally laid out to maximize the axial component of cable strain as compared
with the shear component (Section 2.8.3). If the shear strength and stiffness of the
cablebolt can be increased, then cable strand can be used as shear restraint parallel
to the excavation face (Cutjar et al., 1985; Lappalainen and Antikainen, 1987).
Mandolin style cablebolts installed in this way may also provide tensile capacity
to the underside of a rock beam, similar to reinforcing strand in concrete structures
(Beer and Johnston, 1992; Nickson, 1992).
For example, several cables can
be inserted into steel pipe and
lowered down a drillhole behind
the stope wall as in Figure 2.18.26.
Grout is pumped into the pipe and
around the outside of the pipe. The
cables extend beyond the collar and
are tied back into a second lateral
hole in the wall. This creates a rigid
shear pin which in the case in
Figure 2.18.26 provides shear
reinforcement along a set of sliding
joint surfaces. This method, while
expensive, can eliminate the need
for additional hangingwall
development as required for
conventional cablebolt fans. Note
that without the pipe in this case
the cables alone are unlikely to
have enough shear stiffness to be
effective support.
Mandolin bolting is used in
Figure 2.18.27 in combination with
other patterns to optimize
reinforcement in areas where access
for conventional bolting is limited.
In this case the mandolin cablebolts
provide both shear restraint and
tensile strength. A rock beam with
tensile strength on the excavation
side is stronger than a no-tension
beam (Section 2.18.12).

Figure 2.18.28: Cable sling components
(after Castle and Scott, 1981)
Figure 2.18.27: Cable sling applications
a) U/H cut and fill stopes ; b) roof support
Figure 2.18.28: Drawpoint support
(modified after Stillborg, 1986)
Cable Slings
Another approach to beam building
is the cable sling (Figure 2.18.28)
described by Raju and Ghose (1980),
by Scott and Castle (1981) and Castle
and Scott (1982). The cable sling can
be used to support fill mats and
cemented backfill in underhand cut
and fill stopes (Figure 2.18.29.a) or as
tunnel roof support (Figure 2.18.29.b).
In the Scott system, a cable is inserted
through two opposing Split Sets
(Scott, 1976; 1983). Button anchors
are fitted to the extreme ends of the
cable. Two boreholes are drilled up
and out from the corners of a tunnel
roof and cement is placed in the ends
of the holes. The split sets are pushed
into the holes with the cables,
tightening the intermediate cable
length across bearing blocks to support
the tunnel roof.
In the cut and fill application, the cable slings are installed across the floor of
the current sublevel. Reinforced fabric mats, timber beams and a thin layer of
strong cemented fill is placed on top. After set, the remaining backfill is then
placed on top, and excavation can occur on the next level down, below the slings
and backfill. This technique requires diligence and expertise to be executed safely.
Ore handling systems
Cablebolts can be used to provide
reinforcement and support of drawpoints
and ore handling systems. Drawpoints
normally encompass many of the more
unfavourable conditions for rockmass
integrity as well as plain strand cablebolt
load transfer. They are typically
overstressed during construction and relaxed during service. This reduces the
rockmass stability and is detrimental to the capacity of plain strand cablebolts
(Section 2.6). In addition the vibration and abrasion serve to destabilize existing
and induced structural weaknesses. The design philosophy here, in short, is to use
modified geometry strand, to use plates in backs where possible and to cablebolt
from all accessible directions (Figure 2.18.28). Cablebolt support of orepasses has
also been attempted (Clegg and Hanson, 1992) with limited success.
3 IMPLEMENTATION:
Making the Design Work
3.1 Introduction
The implementation of the cablebolt design involves a variety of mine
personnel working together through a series of tasks. The key steps in the
implementation of the cablebolt system are:
+ Assemble the cablebolting crew. The number of crew members and the skills
required will depend upon the equipment in use and the tasks assigned to the
crew. Establish the payment structure (hourly wage and bonus, or per contract
item), and the contractual conditions for payment of the crew. (Section 3.2).
+ Plan and conduct a training program for the underground personnel involved
in all aspects of the cablebolt operation. Training sessions should be repeated
on a regular basis. The personnel attending the courses should include the
ground control or rock mechanics engineers and technicians, the cablebolting
crews, the drilling crews, the underground supervisors, and the purchasing
department. (Section 3.3).
+ Communicate the cablebolt system design and installation procedure to the
surveyors, drillers, cablebolters, underground supervisors and technicians.
(Section 3.4)
+ Install the cablebolt system, with good control on the quality of the installation.
Any deviations from or problems with the design must be documented and
reported. (Section 3.4)
+ Monitor the quality of the installation procedure, both during the various
installation procedures, and by observing the finished installations. Any
problems with the quality of the installation must be reported and documented.
(Section 3.5).
+ Rectify any problems or difficulties found with the installation procedure, as
soon as they are identified. Problem installations that are likely to compromise
the performance of the cablebolt system should be replaced. (Section 3.5).
The information presented in this chapter has been compiled from a variety of
sources, including observations of practice at underground mines, discussion with

Figure 3.1.1: Key steps in the cablebolt implementation cycle
mine personnel about cablebolt design and installation and a review of the
literature. Every effort has been made to reference published and accessible
information. However much of the detailed information contained in this chapter
can not be referenced in this manner. Therefore, thanks are given here to the
numerous people who provided the practical hints, copies of company installation
procedures and other information upon which this chapter is based.
The procedures and information contained in this chapter have been kept as
general as possible, so as to be applicable to the widest variety of mine sites. The
sample instruction and procedure sheets presented here are intended as guidelines,
that after some modification should be suitable for the specific cablebolt
applications, equipment usage and special circumstances of each site.
Implementation: Making the Design Work 279
3.2 The Cablebolting Crew
The cablebolting crew is the most important component of the cablebolt
installation process. If they do not install the system as designed and the
cablebolts fail, then it is often very difficult to determine the cause of failure: Was
the major cause of the failure poor installation practice or an inadequate cablebolt
design?
There are several factors that influence a cablebolt crew's performance. They
must be well trained, well equipped and supplied, be assigned to the job on an
ongoing basis and work well together as a team. The best examples of crew
conscientiousness and ability come from mine sites where cablebolting is one of
the best paid underground jobs, where the crew have been well trained and stay
with the job for a long time, and where there is frequent communication between
the crew members, the supervisors and the engineers, so that any problems can be
solved in a practical manner as soon as they are identified.
3.2.1 Crew Tasks
The list of operations that may be included in a cablebolt installation is:
1) Locate and orient the hole according to the layout sheets.
2) Drill the hole.
3) Cut the cablebolt strand length specified on the layout, or remove the pre-
cut cablebolts from the palette.
4) Attach spacers at the designated places along the cablebolt strands.
5) Attach or form the end anchor or hanger.
6) Attach the breather and/or grout tubes.
7) Insert the cablebolt(s) into the hole.
8) Form the collar plug, or secure the cablebolt at the hole collar.
9) Thoroughly mix the grout to the specified water:cement ratio.
10) Pump the grout until there is visual proof of return of good quality grout.
11) Seal the end of the breather and grout tubes immediately after grouting.
12) Clean up the grouting equipment and working area.
13) Install the surface fixtures.
14) Report on the installation quality and any problems.
Surveying the hole locations, drilling the boreholes, and installing the
cablebolts are often done by different mine crews. If possible, it is better to have
the cablebolting crew drill the holes, so they are aware of any drilling problems
that have occurred. Adding the responsibility of drilling the holes to the cablebolt
crews' duties may increase the profitability and the appeal of the job as well
(Nickson, 1992).

3.2.2 Crew Composition
The cablebolting crew can consist of one to three people, or more, depending
upon the equipment in use and the installation procedure.
+ A single person can complete the installation of the cablebolts alone when
working with a well equipped, truck mounted, drilling / cable pushing /
grouting system. Some mines have designed and outfitted their own
cablebolting trucks, while others have purchased pre-assembled units such as
the Tamrock Cabolter.
+ A two person crew will be necessary when a scissor lift or fork lift mounted
working platform is provided where the mechanical controls are not
conveniently located on the platform. Two people are also required in
situations where one person is operating the grout pump, while the other grouts
the hole. It is better to have two people placing the cablebolts where long
cables are inserted manually, so that they can share the work load.
+ Three people are required where more than one of the conditions requiring a
two person crew are found. For example, the members of a three person crew
might: 1) operate the mechanical controls for the scissor lift platform, 2) place
the cablebolts and tubes, plug the collars, and grout the holes and 3) assist with
placing the cablebolts and operate the grout mixer and pump.
The crew leader should have the most experience with cablebolting and/or have
demonstrated good problem solving abilities. The leader has the responsibility to
monitor the quality of the cablebolt installation, and to report on any deviations
from or problems with the established procedure. Promotions or transfers of crew
members should be made on a staggered basis so that at least one trained and
experienced person will remain on the crew at all times.
3.2.3 Crew Training
All members of the crew must be taught the correct procedure for all aspects
of the installation procedure, as is discussed in Section 3.3. An important part of
the crew training is to make the crew members aware of the ways in which the
quality of the installation can be compromised, the consequences of quality control
problems and what can be done to improve any problems with the installation.
Crew members must feel free to approach the underground supervisor(s) and
the engineer(s) to discuss any problems with the cablebolting operation. As soon
as any problems are identified, practical and workable solutions must be generated
by discussion between engineering, supervision and the crew.
3.2.4 Crew Payment
The cablebolting crew must be paid at a reasonable level with respect to the
other jobs at the mine. If cablebolting is one of the lowest paid jobs, then the crew
members will request transfer to a better paying, cleaner job, and there will be no
continuity in the crew members' training, leading to poor quality installations.
Crew members should keep a record of reasons for unavoidable down time so
that their pay will not be reduced unfairly. This record should be reviewed
periodically by the supervisors so that any problem areas are identified and solved.
The pay structure for the cablebolting crew at most mines is not linked to the
quality of the job, but is based on the total length of cablebolts installed and
grouted per shift. This policy often results in poor quality installations where time
is saved by mixing the grout continuously instead of in batches, or by pumping the
grout into the holes before it is completely mixed, using dirty cablebolts, by
incompletely plugging the collar, or by not waiting for the return of design grout
consistency along the breather tube or from the collar of a grout tube installation.
If the mine payment structure is such that a bonus is applied to the job, then
consider implementing a system where the bonus is based both on the length of
cablebolts installed and on the quality of the installation. It is important that the
crew not be penalized financially for reporting problems associated with the
installation and quality control of the installation, so ensure that they are well
trained and proficient with the equipment, that the equipment is adequate, and that
all quality control problems have been solved before instituting a quality control
bonus. The quality control checks that can be made are discussed in Section 3.11.
When the cablebolting crew is employed by a contractor, a payment structure
based on detailed quality control specifications should be implemented, and a
quality control checking procedure established and agreed upon. Provision must
be made for the mine engineer to reject any materials or installations that do not
meet the standards established in the contract.
In some South African mines, the payment for support installed is based on the
contracted crew meeting some readily achievable quota of cablebolts and on the
bolts passing quality checks. The quality of the installation is inspected at least
once a day by a Senior Supervisor employed by the mine to monitor quality
control. The supervisor points out any problems to the contractor as soon as they
are observed. The quality is also checked monthly and if the installation does not
meet the level of quality control required, then no payment is made for the work,
and the installation must be brought up to standard at the contractors' expense. A
bonus is paid if the quality of the installation is good. Additional bonus is paid if
the cablebolt quota was exceeded during the month (Thompson, P., 1993, personal
communication).

3.3 Training
Training is an essential component of the cablebolt implementation process.
The training course should educate everyone involved in the design and the
implementation of the cablebolt system so that they understand the function of the
cablebolts, the proper procedures for all of the equipment in use, and the
consequences of poor quality control. It is also important to establish the
responsibility of each department for each step of the implementation process.
The information to be imparted during the training sessions can be structured
by answering the questions Why?, What? and How? with regard to the
cablebolting process. Methods for providing feedback on the installation process
and safety issues must also be discussed during the course.
A suggested table of contents for the training course is given in Table 3.3.1.
The table makes reference to the sections in this book where further information
on each subject in the training course can be found.
The training material must be updated as new techniques and products become
available, if problems with installation of regularly observed or as company
policies are altered. The training program should be simple and easy to conduct
so that it can be done at any time that it is thought necessary.
Anyone who is involved with designing, purchasing materials for, supervising,
checking or implementing the cablebolt system should attend the training sessions.
There may be as many as three different underground crews working on the
cablebolt installation on each shift, including the surveying crew, the drilling crew
and the cablebolt installation crew. All members of these crews should attend the
courses.
Training courses should be conducted frequently. The entire course should be
taught annually so that the information discussed during the course remains fresh
in the minds of the personnel involved with the cablebolting process. Periodic
repetition of the training courses also provides the opportunity for review and
discussion of the design and implementation of the cablebolt system. Whenever
new people join the cablebolting crew, or if the performance of the crew appears
to be declining, the appropriate sections of the course should be repeated.
A break in period should be established after the first course has finished to
give the crew time to become proficient with the equipment and procedures.
During this period of time, the engineers, technicians and supervisors should visit
the crew more often than usual to observe their work and to solve any problems
with the installation procedure that might arise.
Whenever different materials or equipment are to be used in the cablebolt
operation, the engineer should work with the crew to check that the new items will
perform as expected, and to sort out any problems with the installation procedure.
Copies of relevant drawings, tables and procedural instructions should be
provided to course attendees and be regularly updated as procedures or equipment
change.
Table 3.3.1: Suggested table of contents for the cablebolt training course.
Module 1: Why use cablebolts? Section 3.3.1
1) Purpose of cablebolting.
2) Application(s) of cablebolting at the mine.
3) Function of cablebolts
Module 2: What is a cablebolt? Section 3.3.2
1) What is a cablebolt?
2) The components of a cablebolt element.
3) The optimum cablebolt element for the site.
4) How does a cablebolt work?
Module 3: How are cablebolts installed and checked? Section 3.3.3
Classroom discussion:
1) The cablebolting cycle, with emphasis on where the implementation process
fits into the cycle.
2) The steps in the implementation process.
3) Overview of the cablebolt installation method(s).
4) Cablebolt layout and design specification sheets.
5) Implementation procedures and safety.
6) Quality control guidelines.
7) Reporting on the cablebolt installation process.
8) Monitoring the cablebolt installation process.
Hands on training:
1) Purchasing and handling the materials.
2) Installing the cablebolts using layout, procedure and observation sheets.
Module 4: Safety. Section 3.3.4
1) Review and discussion of safety issues.
Module 5: Feedback on the installation procedure. Section 3.3.5
1) Review and discussion of feedback recorded on the layout sheets.
2) Review and discussion of feedback recorded on the observation sheets.
3) Review and discussion of the quality control check list.

3.3.1 Why Use Cablebolts?
The intent of this module is to explain in general why support is used at the site
and in particular why cablebolts are used in certain environments. This discussion
should be very specific to the mining environment and also to the trainees' level
of knowledge. Some points that should be included in the discussion are:
1) Purpose of cablebolting. This may include such items as improving the
safety of workplaces and the stability of the stope boundaries as well as
reducing dilution and oversize (Section 1.2).
2) Application(s) of cablebolts, including a general discussion (Section 1.3),
site specific support patterns and the intended support mechanism of the
cablebolt systems in use at the mine site.
3) An introduction to cablebolt functions (Section 1.6) is useful to avoid some
on-site adjustments to procedure which may impair support effectiveness.
3.3.2 What is a Cablebolt?
This module describes a cablebolt, including:
1) A description of a typical cablebolt (Section 1.1), including the basic steel
strand, the cablebolt toolbox (Section 1.5), grout and any surface fixtures.
Detail can be added from the information presented in Chapter 2.
2) An introduction to the components of the cablebolt element in use at the site.
The cablebolt element includes the cablebolt strand(s), the installation
components such as hangers and tubes, the grout mix and any surface
fixtures. Describe all of the combinations of the element components that
will be used in different applications at the mine.
3) A description of the optimum installed cablebolt element. This will depend
on numerous factors and can only be fully developed once the cablebolt
system has been designed and selection of the best implementation
procedure has been made. As an example though, an optimum cablebolt
element could be a clean cablebolt installed at the exact position designed,
in the centre of the hole and with a full column of completely mixed grout
of the correct water:cement ratio. The optimum cablebolt could also include
a clean plate of the design dimensions securely fastened with a clean,
matched barrel and wedge anchor to the end of the correctly tensioned clean
cablebolt strand. Some suggestions for the conditions for optimum
installation are given in the left hand column of the quality control guideline
tables in Sections 3.8.3 and 3.9.3. A discussion of the items which are
applicable to your site and usage of cablebolts would provide useful
information for the crew and supervisors.
4) A discussion of how a cablebolt works, including the interaction of plain,
multiple and modified strand with the surrounding grout. Include a brief
discussion on the impact of stress change (Section 2.6).
3.3.3 How are Cablebolts Installed and Checked?
The cablebolt implementation module should include a general overview of the
implementation process, a detailed discussion of the implementation steps and
associated instruction documents, and a hands on training session.
Samples of the sheets used in this module are given in Sections 3.7 to 3.9:
Design specifications: Material purchasing
Cablebolt layout: Plan and Section
Cablebolt installation: Layout and Notes
Procedure and safety: Material handling
Borehole drilling
Cablebolt placement
Grout mixing and pumping
Surface fixture installation
Quality control guidelines: Cablebolt borehole preparation
Cablebolt placement
Grout mixing and pumping
Surface fixture installation
Feedback: Drilling observation report
Cablebolt installation observation report
Cablebolt quality control check list
The classroom session should include discussion of:
1) The need for effective communication throughout all aspects of the work
(Section 3.4).
2) How the implementation process fits into the cablebolting cycle. This
discussion should provide the trainees with an overview of the cablebolting
cycle (Section 1.4), including general information about the design and
verification processes, so that they will understand the need for feedback.
3) The steps in the installation process (Section 3.6). Everyone involved in the
process should know where they fit in to the whole picture.
4) The installation method(s) that will be used at the site. Discuss the different
installation methods (Section 1.8), and the advantages and disadvantages of
each.
5) Describe the range of equipment available on the site. A sample list is
provided in Section 1.7.

6) The cablebolt layout and design specifications: Sections 3.4, 3.6, 3.7.1,
3.8.1, and 3.9.1. Work through the sample sheets during the course to make
sure that everyone understands the information contained in the sheets and
when feedback is required.
7) Implementation procedures and safety, modified from samples provided in
this book and modified to reflect conditions at the site: Sections 3.6, 3.7.2,
3.8.2, and 3.9.2. Discuss the contents of the sheets briefly, emphasizing the
allowable tolerances for deviation. This information can also be reviewed
and is easiest to teach in detail during the hands on training session.
8) Quality control guidelines: Sections 3.5, 3.7.3, 3.8.3 and 3.9.3. This
discussion must emphasize that the cablebolt will only perform as designed
if the installation quality control is good. Discuss the correct procedures for
good quality control and solutions for commonly encountered quality
control problems.
9) Reporting on the cablebolt installation process: Sections 3.4, 3.5, 3.6, 3.7.4,
3.8.4 and 3.9.4. Discuss the observation sheets that are to be completed by
the crews, and the possible uses of the information recorded on these sheets.
10) Monitoring the cablebolt implementation process: Sections 3.5 and 3.11.
Describe the methods used for monitoring the grout quality during
installation and the check lists that the supervisors will complete after each
visit to the work site.
The hands on training session can be conducted in a working area that is to be
cablebolted, or anywhere else using pipe pumping tests. In the latter, a length of
pipe simulates the borehole. One of the advantages of the pipe pumping tests is
that the pipes can be cut apart after the tests to check the completeness of the grout
column. A description of the materials required for and the procedure to be
followed during a pipe pumping test is given in Chapter 2.
The installation trials provide the opportunity to catch quality control problems
before they become established installation procedures. In addition, it will be
apparent during the trials whether the equipment and materials are adequate for the
specified design and work site conditions. The hands on training sessions should
expose all of the people involved in the implementation process to:
1) Selection and handling of materials following the procedures discussed in
Section 3.7. Collect samples of the different materials available and create
examples of well stored and handled materials and of materials that should
be rejected due to deterioration of their quality. It is instructive for everyone
involved in the cablebolt implementation process to see the differences
between the well stored and poorly stored samples.
2) Installation of cablebolts using the sample layout, procedure and observation
sheets that were discussed in the classroom sessions. Everyone should
complete the feedback sections on the layout and specification sheets, the
observation reports and the quality control check lists.
3.3.4 Safety
Safety pointers are included in the procedures but should be discussed during
the installation training as well. A list of some safety points is given in Section
3.6.1. Review the safety guidelines after the hands on training to reinforce the
ideas, and to ask the trainees for additional safety suggestions. Add useful
suggestions to the list and to the procedure instructions.
3.3.5 Feedback on Installation Procedures
Feedback is essential, and this module reviews how it is recorded and explains
how the reports are used. The following points of discussion should be covered
during the training course:
1) Daily feedback. The crew is required to provide feedback on a daily basis
about the drilling and installation process on the layout sheets, including any
deviations from design due to insurmountable problems, and usage of
cablebolts, cement, and other materials. Notes of layout changes must be
recorded so that they can be considered when back-analyzing the
performance of a cablebolt layout or system design. If layout changes occur
frequently, then the reason for the changes should be determined and the
problem solved. For example, if the grout rarely returns down the breather
tube, then the diameter of the breather tube, the grout fluidity, and the pump
capacity should all be examined to try to determine the source of the
problem.
2) Periodic feedback. The underground supervisors should also complete a
check list for quality control issues after they have observed the crew(s)
working underground. A sample quality control check list is given in
Section 3.11 an should be modified to reflect the specific situation at the site
prior to use in the training session at the working face. The information
recorded on the check list should be used to determine if there are any
recurrent problems in the installation process which need to be worked on,
and where applicable, to calculate the payment of bonus to the crews.
It would be instructive to work through an example of how data recorded on
feedback or observation forms could be used by the engineer. After the
introduction of these forms, as everyone at the site becomes familiar with the
cablebolting cycle, some examples from the site should be incorporated into the
course material. In the meanwhile though, the case history of quality control
improvement given in Section 3.12 could be given to show the utility of
communication and feedback. In the case example, the quality control problems
only came to light when large wedges fell from the drift back. Had the drilling.
installation and quality control guidelines been in place prior to the failures, the
mine personnel might have been alerted to the problem sooner.

3.4 Communication
Communication between all people involved in the cablebolt implementation
process is important to ensure that the operation runs smoothly and efficiently, and
that any problems are resolved as soon as they arise.
There are two levels of communication that should occur: regular and periodic.
Regular communication should include:
1) The engineer communicating the design layout and specifications to the
drilling and installation crews: Layout and specifications.
2) The engineer instructing the crews in the procedures to be followed during
the implementation process. Safety issues must be included in these
instructions: Procedures and safety guidelines.
3) The engineer and supervisors informing the crews of the potential quality
control problems and their solutions: Quality control guidelines.
4) The crews providing daily feedback on the installation process, including
usage of materials, any problems with equipment and materials and any
deviations from the design as specified on the layout: Feedback.
Periodic communication involves:
1) The supervisors, engineers or technicians recording their observations of the
quality of the installation during spot checks: Quality control check list. This
information should be reviewed frequently by the engineer.
2) The engineer discussing, with the crew, any problems that have been
identified with the installation process . The presence of problems could be
indicated by the comments recorded on the feedback forms, on the Quality
control check list or by the results of grout tests.
3) All personnel involved in the cablebolt cycle discussing the design,
installation procedure and verification of the process during monthly
meetings. The monthly meetings should provide a forum for the transfer of
information and discussion on the installation method, equipment and
materials. A suggested agendum for the monthly meetings is:
+ General review of operations for the month, including the total length of
cablebolts installed, crew productivity and materials consumed. These items
should be compared to the budgeted work.
+ A review of the work planned for the next month.
+ Review of post-installation quality control checks. Resolutions must be
found for any quality control problems found during these checks.
+ Discussion of any quality control problems reported by the crew, supervisors
or engineers on the feedback reports or on the quality control check list.
Review of the solutions adopted and discussion of other possible solutions
that might also work.
+ Discussion of any procedural problems, and possible solutions.
+ Description of any changes to the installation procedure.
+ Suggestions for improvement to the cablebolt system design or installation
procedure.
+ Discussion of any new products under review or available: advantages and
disadvantages.
A spreadsheet listing the annual cablebolt budget, which details the metres of
cablebolts required for each planned stope, should be produced and available to
crew members (Boaro, J., 1993, personal communication). This spreadsheet
would aid cablebolt planning and progress review.
The communication of feedback is essential so that any problems with the
installation equipment or any suggestions for improvements to the cablebolt design
or installation procedure will be considered as soon as they arise. Changes to the
cablebolt layout due to problems encountered during the installation process
should be recorded for back analysis of any subsequent problems with support
performance. If it is felt that the installation problems are severe enough to
compromise the capacity and designed function of the cablebolts, then additional
cablebolts should be designed and installed in the problem area.
Copies of the original design specification sheets, and the feedback comments
must be kept in a readily accessible place. Then if a rockmass or support system
failure occurs, a description of any problems encountered during the installation
will be available as basic information for the evaluation of the cause of the failure.
A failure report documenting the volume of and suspected reason for the failure,
assessing the performance of the cablebolts, and discussing anticipated influences
of the failure on adjacent unmined stopes should also be filed.
The crew productivity can be assessed periodically from the information
recorded on the reporting sheet. Reasons for drops in productivity should be
determined and corrected, and most importantly, suggestions for productivity
enhancing changes evaluated and implemented quickly.

3.5 Quality Control Practice
The cablebolts must be installed with the best quality control possible. The
designed function of the cablebolts can be severely reduced if the quality of the
installation is poor, as is discussed in further detail in Section 3.11. The
maintenance of good quality control during installation is the responsibility of
everyone involved in the implementation process:
1) The crews should be encouraged to monitor the quality of the materials,
equipment and the installation process, and report any problems as soon as they
arise even if the crew has solved the problems during the shift. The Quality
control guidelines presented in this handbook are intended to be given to the
crews by the engineer along with the design specifications and procedures. The
guidelines list good practice procedures and solutions for some installation
problems. Samples are given in Sections 3.8.3: Cablebolt borehole preparation
and 3.9.3: Cablebolt placement, Grout mixing, Grout pumping, and Surface
fixture installation. The guidelines should be reviewed by the crew
periodically, and at any time when the installation procedures change.
Additions and improvements to the guidelines are encouraged as merited by
unique site conditions or by experience and procedural improvements.
2) The underground supervisors, technicians and engineers should also monitor
and report on the quality of the materials, equipment and the installed
cablebolts during spot check visits to the crews. Observations of the
installation process should be recorded on the Quality control check list
(Section 3.11). Solutions to any problems should be found as soon as possible
in consultation with the crew and implemented immediately.
When quality control problems are found, their impact on the cablebolt system
function should be evaluated as discussed in Section 3.11. If the system capacity
is severely compromised by poor quality control, then additional cablebolts may
have to installed in the working area.
Where the payment of bonus to the crews is based in part on the quality of the
installed cablebolts, reported, unavoidable problems should not be penalized by
reduction of the bonus payment. Solutions to these problems for future
installations should be investigated and implemented as soon as possible.
New equipment and hardware is constantly being developed and supplied to the
mining industry. These improvements are welcome but greet with skepticism any
claims that quality control is no longer required. Old quality issues may become
less critical but new ones will arise. It is better to maintain a high standard of
quality regardless of the technology in place. Diligence costs little and always
brings rewards.
3.6 Installation
The cablebolt implementation process involves several steps as shown in Figure
3.6.1. Instructions for each of these steps are given in the procedural guidelines
presented in the following pages. Procedures are also discussed by Hunt and
Askew (1977) and Schmuck (1979).
The guidelines presented here have been kept general, in an attempt to cover
all aspects of the procedure for each of the installation methods presented in
Chapter 2. Create the following set of procedural guidelines and feedback report
forms, specifically for your site and use of cablebolts and equipment:
1) Material purchasing specification.
2) Material handling procedure.
3) Borehole drilling procedure.
4) Drilling observation report.
5) Cablebolt placement procedure.
6) Grout mixing and pumping procedure.
7) Surface fixture installation procedure.
8) Cablebolt installation observation report.
In the sample procedures, a number of entries have been printed in italics.
These and any other items should be changed to reflect the practice and procedures
appropriate to your usage of cablebolts before the sheets redistributed.
In any installation, problems may arise at the working face, such as not being
able to set up the drill rig in the exact position required by the design. In the
general procedure guidelines, an attempt has been made to give some allowable
tolerances for deviations from design. Then if problems do arise, the crew can
alter the design within the specified limits at the working face, and will not be
delayed. Allowable tolerances have been included in the following guidelines in
italicized text. When creating the set of procedural instructions for the crew(s),
alter these tolerances to reflect the conditions of your site. The specific instructions
for the installation of each cablebolt are presented on the following design
specification sheets.
1) Cablebolt layout plan.
2) Cablebolt layout section.
3) Cablebolt installation layout specification and notes.
These sheets provide all of the information required by the drilling and
cablebolting crews to completely install the cablebolts designed for a particular
working area. These sheets also provide opportunities for the crew(s) to provide
feedback on the installation process.

Design Specification and
Procedures
Quality Control
Create layout sheets for each
cablebolt ring. Record all of the
installation instructions on procedure
sheets for each step of the process
given below.
For each step in the
implementation process,
monitor the quality and
solve any problems. Make
changes to the specs. and
procedures as required
Material purchasing,
storage and transport
Borehole surveying and
drilling
Ensure conformance with
material specifications,
and reject if substandard.
Keep materials dry, clean
and at moderate
temperature.
Survey and mark the
collar positions of the
holes. Drill the hole in the
position and at the angle
specified. Clean the hole
by flushing with water.
Cablebolt placement
Attach spacers, hangers and tubes to
the cablebolt. Ensure that the
cablebolt is secure in the hole and that
the hole collar is well sealed.
Grouting
Mix the grout to the specified W:C in a
batch, ensuring that it is well mixed.
Grout the hole, and tie off the tube(s).
Tensioning and Plating
Ensure that the plate is flush with and
the cablebolt perpendicular to the rock
surface. The correct barrel and wedge
combination must be used, and the full
tension applied to the cablebolt.
Figure 3.6.1: Steps in the implementation process.
Implementation
3.6.1 Safety Guidelines
There are numerous safety issues that should be considered during the
installation of cablebolts, which must be stressed during the training course. The
following guidelines provide a partial list only, and must be expanded to cover all
aspects of safety at the site. The safety guidelines should be repeated in the
appropriate procedures to stress their importance. Review these guidelines
regularly to ensure that they are up to date.
+ Ensure adequate ventilation is turned on.
+ Collect all required supplies, tools and safety equipment.
+ Review all hand and light signals with new crew members.
+ Make the working area safe by scaling down loose rock.
+ Clear the floor of the working area, to allow for good footing.
+ Store tools, materials and equipment in safe, out of the way places to avoid
tripping hazards.
+ Use a certified safety restraint system (e.g. rope, lanyard or full body harness)
when working on platforms.
+ Stand in the middle of the roll of cablebolts when cutting the packing straps.
Cut the straps in the correct order to allow the cablebolt strand to unroll slowly,
in a controlled manner. Ensure that all other crew members are far enough
away to be out of the range of any springing or lashing cablebolts.
+ Wear a face shield, safety glasses and work gloves when cutting cablebolts with
an abrasive cutter, since small shards of metal and sparks are created.
+ Adopt safe-lifting practices when pushing cablebolts into holes. Guide the
free end of the cablebolt element as it is being inserted into the hole so that it
doesn't whip about.
+ Do not stand directly beneath holes, especially when sealing the collar with
grout or resin or when grouting the hole.
+ Ungrouted cablebolts can fall out of upholes, with the potential to injure people
working below. Post a warning sign, which indicates that there are ungrouted
cablebolts overhead, at all entrances to the work area.

+ Grout can severely burn exposed skin. Wear glasses, gloves and protective
clothing when working with cement powder and grout. Barrier cream should
be used on any exposed skin, and on the hands. If any grout does touch bare
skin, the area should be washed immediately with soap and water and
completely rinsed. If grout gets into your eye, flush with fresh water, and
immediately report to the First Aid station.
+ Ensure that water and air supplies are turned off while assembling and
disassembling the mixer and pump.
+ All clothing and hands must be clear of the mixer paddles before starting the
mixer motor.
+ Wear a dust mask when breaking open and handling bags of cement.
+ Ensure that the grout and breather tubes are of the correct pressure rating, and
have not been damaged in any way.
+ The pressure created during the grouting operation has the potential to burst the
grout hoses or tubes. Turn the pump on slowly and keep the pumping rate at
a slow speed to prevent excess pressure build up. On the other hand, do not
reduce the pressure too much, or else the pump will stall.
+ Use caution when disconnecting hoses and tubes, because pressurized grout
may still be present in them.
+ Clean up all empty cement bags and cotton waste, as they are a fire hazard.
The safety clothing and equipment provided to the cablebolting crews should
include:
+ Oiler or other tough waterproof pants.
+ Long water proof gloves and barrier cream.
+ Safety glasses, full face shield or protective goggles, and antifog solution.
+ Dust masks.
+ Hearing protection.
+ General safety equipment including work gloves, safety ropes, lanyards, well
fitting safety belt and correctly sized full body harness.
3.7 Material Purchasing and Handling
The equipment required for the cablebolting job, such as grout mixers and
pumps or tensioning jacks, is discussed in Chapter 2, and is generally purchased
once every few years. This section covers the materials that are consumed on a
regular basis by the cablebolting operation.
The general list of materials required for cablebolt installation should be
specified by the engineer so that the crews can ensure that they have all of the
materials required for a particular job. A sample list of required materials given
in Section 3.7.1 is fairly general, and should be discussed with the crew to ensure
that it is complete.
Quality Control
As the materials arrive on the site, they should be checked for quality, and
rejected if the quality is poor, as discussed in Section 3.7.2. The materials are
generally stored in a cool dry location on surface, and transported underground as
required. The storage period at the mine site should be minimized by ordering
smaller quantities of materials more frequently, to reduce the chance of
deterioration of the quality of the material. All people working on the storage,
transport and handling of the materials must endeavour to keep them as clean and
dry as possible.
The quality of the materials purchased is very important as well. All materials
must meet or exceed the specifications supplied by the engineer. For example, the
cheapest cement or cablebolt strand may not have the performance characteristics
required for the design. Where cost savings appear to be substantial for an
alternate material, the purchasing department should consult with the engineer
before trying a new product. If the specifications of the product are not proven,
then field trials of the product should be arranged and conducted prior to volume
purchases.
Feedback
Lapses in quality of newly shipped materials should immediately be brought
to the attention of the supplier so that the problem cab be dealt with immediately.
It is advisable for the purchasing department or the engineer to keep abreast of
the new development of new materials or equipment. The last few years have seen
great changes in the cablebolting market. As more experience is gained and as
more mines use cablebolts in different environments, manufacturers of
cablebolting hardware and equipment can implement improvements based on this
experience. Feedback from the mines to the suppliers and manufacturers should
result in continued product development and refinement.

3.7.1 Design Specifications
Specification: Material purchasing (Example)
Stope:
Material storage location:
Date materials required at the storage location:
General:
* Axe, hammer, hacksaw, vice grips, crescent wrench, pipe wrench, electrical or
duct tape, knife, air powered cutter or grinder with blades, string, tape measure.
Hole drilling:
#
#
#
*
*
*
Long hole drill.
3' to 4' drill rods, rigid couplings and one centralizing coupling.
Steel rod rack.
Sample bags for drill cuttings.
2 1/2 diameter drill bits.
Air and water hoses.
Placement: (This example is for a breather tube installation of 300 cablebolts, with a
10% surplus contingency, and using Procedures A1, B3, D2, E3, F1, G1, H1, I1, J1,
K2 and L1 in Section 3.9.)
330 pre-cut plain strand cablebolts of 10 m length, with nut for hanger attached.
660 m of 19 mm I.D. 250 psi pressure rated grout tube.
4300 m of 13 mm I.D. 100 psi pressure rated breather tube.
990 7.5 cm by 2 cm spring steel clips.
330 4.5 cm diameter washers.
330 wooden wedges.
Cotton waste.
Grouting:
*
*
*
Spedel 6000 grout mixer and 3100 pump.
2 x 20 litre clean, empty pail.
2 cement scoops.
35 grout tube connectors.
660 25 kg bags of Portland Type 10 cement, supplied on wooden palettes
wrapped in plastic.
Surface fixture installation:
*
330 20 cm by 20 cm by 1 cm flat plates.
330 wedge and barrel sets.
Tensioning jack.
* Note that the items marked with a * are purchased periodically and will not normally
be specified for purchasing on a stope by stope basis. In addition, these items are
generally kept with the crew at all times. It is the crew members' responsibility to
request from the purchasing department these materials when they are required.
# The items marked with a # are purchased once every few years and are generally
with the crew at all times.
3.7.2 Procedure and Safety
Procedure: Material handling
Procedure:
1 Order the materials listed on the purchasing specification sheet. If the exact
materials specified are not available, discuss the alternatives with the engineer.
2 Check the materials on arrival at the mine site for quality. Reject any shipments
of sub-standard materials and record the reason. Standards are listed below.
3 Store the materials in a dry, clean, shaded area which remains at a moderate
temperature (between 20 and 40-C). Do not stock pile the materials for too
long, otherwise their quality will start to deteriorate.
4 Deliver the materials required for each working area at least one shift before the
start of work in that area.
5 If the crews report any problems with the quality or ease of use of the materials.
This should be discussed with the engineer.
Material quality standards:
Cablebolt strand:
-
-
-
Clean and free of rust and grease or oil.
No nicks or surface damage to the steel strand.
Coils no smaller than the minimum specified diameter.
Cement:
-
-
-
-
Stacked on a wooden palette which is well water proofed with plastic wrapping.
Do not stack more than 2 palettes high.
Lump free.
Do not leave in the sun for long.
Placement materials:
-
-
Grout and breather tubes must arrive well secured in a roll, be clean and have
no kinks or bends.
Fabric for collar plugging must be soft, and not too stiff.
Surface fixtures:
-
-
Plates, wedges and barrels clean and free of rust.
The wedges and barrels must match - when placed over a piece of cablebolt,
the narrow ends of the wedges must not protrude through the end of the barrel,
and the tops of the wedges must be at least 10 mm above the top of the barrel.
Please make a note of the date and supplier of any materials that are rejected due to
sub-standard quality and the reason for the rejection.

3.8 Cablebolt Borehole Preparation
Sample versions of the Cablebolt borehole preparation sheets for:
Design specification: Cablebolt layout plan and sections.
Procedure and safety: Borehole drilling procedure.
Quality control: Cablebolt borehole preparation quality control
guideline.
Feedback: Drilling observation report.
are given in the following text. These sheets are samples only, and should be
modified to suit the particular cablebolt application, specific drilling equipment
in use, and safety concerns at your mine. At some sites, several versions of these
sheets will be required to cover all possible combinations of cablebolt application
and installation procedure.
The Cablebolt layout design specification presents the information required for
locating the cablebolt holes. This sheet is used by the survey crew when locating
and marking the collar position of the boreholes on the rock face. The surveyors
will either mark the collar location of every hole, or just the middle of each
cablebolt ring, depending upon the accepted practice at the site. If there are any
problems with locating the hole collar positions as designed, the surveyors should
note on the layout sheet the new distance between the reference point and the
centreline ($) of the ring.
The cablebolt boreholes are then drilled by the drilling or cablebolting crew at
the angle and to the length detailed on the Cablebolt layout sheet, following the
instructions given on the Borehole drilling procedure sheet. If the crew have any
problems with drilling the holes at the specified pattern, the layout can be altered
within the allowable deviations given in the procedure. Changes that are made to
the collar location or the borehole length or inclination should be noted and drawn
on the Cablebolt layout section sheet. Changes to the position of the cablebolt
ring and the date on which each ring of boreholes was drilled should be recorded
on the Cablebolt layout plan sheet.
During borehole drilling the crew should monitor the following items:
1) Loss of drilling water.
Drilling water loss indicates the presence of an open structure such as a
joint, fault or void. Where there are open joints, there may be problems with
grouting the borehole, if the grout flows away into the rockmass. Section
3.9.2: J3 to J5 describes techniques for grouting in jointed rock. In addition,
open structures or voids can have a great deal of influence on the rockmass
behaviour. Reports by the drilling crew of such features must be evaluated by
the engineer to determine if unexpected rockmass failure could occur.
2) Change in drilling rate.
An accelerating drilling rate indicates that softer rock has been encountered.
Softer rock provides less confinement for the cablebolt, reducing the bond
strength of the cablebolt. A soft, easily drilled layer may also indicate the
presence of a fault zone, which could lead to unanticipated rockmass failure.
A decelerating drilling rate will occur when harder rock is encountered, but is
generally of no concern.
3) Changes in the appearance of the cuttings.
The appearance of the cuttings should be uniform for a hole drilled in a
single lithology rockmass. The expected rock types along the drill holes should
be shown on the Cablebolt layout section sheets. Then if the cuttings are
different than expected, the drillers should describe the cuttings and take a
sample of the cuttings for later inspection. If the cuttings indicate a much softer
or weaker zone of rock than was considered during the design of the cablebolt,
then the capacity of the cablebolt system will be reduced. It may be important
to assess the influence of different lithologies on the cablebolt bond strength.
4) Holes producing water.
Excess water in the hole may make grouting difficult, and is likely to mix
with the grout, increasing the water:cement ratio (W:C) above design levels.
Assess the possible increase in the W:C of the installed grout. If the increase
in W:C is likely to be significant, reduce the amount of water in the mix design
to account for the presence of water in the borehole.
Any changes in these items should be recorded on the Drilling observation
report. The engineer must review the report periodically and determine if any of
the observations recorded indicate unexpected rockmass failure modes, or
problems with the cablebolt installation or with the expected cable performance.
A discussion of the quality control problems that can be encountered when
preparing the borehole is given in the Cablebolt borehole preparation quality
control guidelines. There may be several additional quality control issues that
should be added to the guidelines produced for each site. It is important that the
drilling crew review these guidelines frequently until they are fully aware of the
contents, so that they will be able to assess the influence of problems and
understand the importance of reporting on quality control problems.

3.8.1 Design specifications
Cablebolt layout: PLAN (Example)
Plan #: 1450/-1 Uphole Cablebolts
Stope #: 1400/C Cablebolt ring #s: A to S
Reference point (RP): #: 1450 - RP3 Location: N: 1078 E: 2430 Elev.: 1450
Reference line: Azimuth: 85/0
Cablebolt
ring
Specified
Distance:
RP - ring
Feedback
Surveying * Drilling
Distance:
RP - ring
Date Crew
Leader
Distance:
RP - ring
Crew
Leader
A 12.5
B 15
C 17.5
D 20
E 22.5
....
* Mark the reference line and ring position on the back or walls of the drift.
Design Specifications: 3.8.1 continued
Cablebolt layout: SECTION (Example)
Stope #: 1400/C Plan #: 1450/-1
Cablebolt ring: A
Drilling crew leader: Date holes drilled:
Cable-
bolt #
Specification: Feedback:
Dist.
from
$ (m)
Hole
Diam
(mm)
Dip
-
Dump
-
Length
(m)
Dist.
from
$
Dip Dump Length
1 -2.8 54 -36 0 11
2 -2.3 -27
3 -1.6 -18
4 -0.9 -9
5 0 0
6 +0.9 +9
7 +1.9 +18

Procedure: Borehole drilling
Procedure:
1 Make the working area safe by scaling down any loose rock. Clear the floor of
the working area of any debris to allow for good footing.
2 Collar each hole as close as possible to the location specified on the Cablebolt
layout plan and sections. Definitions of the terms Dip and Dump are given at
the bottom of this sheet. If there are hole obstructions, remove them; for
example trim the screen, scale the loose or move the pipes. Don't drill through
any steel hardware, such as straps or plates.
3 Select the drill bit for the hole diameter specified. Make sure it is not too worn.
4 Drill each hole at the angle and to the length specified, keeping hole wander to a
minimum.
5 Sample the drill cuttings for the holes as specified, and at any time that the
cuttings look different than expected.
6 Blow the borehole clean, with compressed air, as soon as you have finished
drilling. If necessary, flush the hole with water to ensure that it is clean.
7 Block the collar of any downholes against the inflow of mud or dirt.
Allowable deviations and Feedback:
If there are problems, and it is impossible to drill the holes as specified:
-
-
-
the position of the collar may be moved by up to 100 mm,
the borehole angle can deviate by ± 2.5-, and
the length of the borehole can be exceeded by 1 metre, but can't be too short.
Record any changes to the layout on the Cablebolt layout plan and sections.
Observations and Feedback:
If you observe any of the following changes while drilling the boreholes, please note
them on the Drilling observation report, indicating the position of the changed zone:
-
-
-
-
Loss of drilling water.
Drilling rate faster or slower than usual - estimate the difference in the rate.
Hole cuttings different than usual - describe the appearance of the cuttings.
Holes producing water - estimate the volume of water in litres / second.
Definitions:
3.8.3 Quality Control
Quality control guidelines: Cablebolt borehole preparation
Quality control Consequences of poor
quality control
How to achieve good quality
control
Good Poor
Hole size
correct
Hole size
too large
or too
small
If the borehole is too large,
additional grout will have to
be pumped into the hole,
and the spacers and hanger
will be too small. If the hole
is too small, it will be difficult
to insert the cablebolt,
spacers and hanger into the
borehole.
Drill the borehole to the
specified diameter, and
ensure that the bits in use
are not too worn.
Correct
hole
length
Hole too
long or
too short
If the borehole is too long,
then extra grout will have to
be pumped into the hole.
If it is too short, then the full
length of the cablebolt will
not fit into the hole. This
could result in flared sections
of the cablebolt being left
hanging out of the collar of
the borehole creating plating
problems.
Drill the borehole to the
specified length. Report any
boreholes that are drilled to
a length different from that
specified.
Straight
drill hole
Hogged
drill hole
The cablebolt will be in
contact with the borehole
wall in some places over the
length, and so the
circumference will not be
completely embedded in
grout. This may result in
reduced capacity of the
cablebolt. The toe end of
the cablebolts will not be in
the designed positions, which
may lead to failure of the
cablebolt system.
Drill the borehole as straight
as possible.
Borehole
flushed
clean
Borehole
not clean
Any mud or cuttings
remaining in the borehole
will reduce the length or
diameter of the grouted
column. Any water left in the
hole prior to grouting is likely
to dilute the grout. Any oil
left in the hole will reduce the
cablebolt bond strength.
Flush the borehole clean
with water and keep it clean
by plugging or capping the
collar of downholes until
ready to install the cablebolts
and grout the borehole.

3.8.4 Feedback
Feedback: Drilling observation report
Plan #:
Stope #: Cablebolt ring #s:
Date: Drilling crew Leader:
Cable
ring #
Cable
hole #
Drilling
water
loss
Change in drilling rate and/or
Change in appearance of cuttings:
Distance from collar (m) Description
Bore-
hole
water
@ (m) @ (m) Description (l/min)
Drilling rate: Estimate the change: e.g. 2+ = 2 x faster, 2- = 2 x slower
Appearance of cuttings: Describe the cuttings.
Any changes to the borehole layout must be recorded on the Cablebolt layout
plan and section specification sheets.
3.9 Cablebolt Installation
The sheets required to fully communicate the cablebolt design are:
Design specification: Cablebolt layout plan and sections.
Cablebolt installation layout specifications and notes.
Procedure and safety: Cablebolt installation procedure, including
Cablebolt placement, Grout mixing, Grout pumping,
and Surface fixture installation.
Quality control
guidelines:
Cablebolt placement.
Grout mixing.
Grout pumping.
Surface fixture installation.
Feedback: Cablebolt installation observation report.
A sample version of each of these Cablebolt installation layout sheets is given
in the following text. These sheets are samples only, and should be modified to
suit the particular cablebolt application, specific drilling equipment in use,
grouting method selected, safety concerns and allowable deviation tolerances at
your mine. The Cablebolt installation layout specification sheet provides all of the
information required to implement the cablebolt design in the pre-drilled
boreholes. A layout sheet should be prepared for each ring of cablebolts to allow
the crew to record feedback about each cablebolt installation. At some sites,
several versions of these sheets will be required to cover all possible combinations
of cablebolt application and installation procedure. In addition, all of the columns
on the sheet may not be required, for example, where only one type and length of
cable are used these sheets could be simplified. Wherever possible, the safety
guidelines listed in Section 3.6.1 have been repeated here to emphasize their
importance on a daily basis to the crews.
The installation instructions are detailed for the crew on the Cablebolt
installation procedure sheet, which includes information about the cablebolt
placement, grout mixing, grout pumping and surface fixture installation.
Quality control problems that could be encountered during the installation of
the cablebolts and their solutions are listed in a set of guidelines for each of the
steps in the installation process in Section 3.9.3. These guidelines are by no means
complete, and should be added to if quality control problems arise at your site.
Feedback on the installation process is made by the crew on the Cablebolt
installation layout sheet and on the Cablebolt installation observation report.

Cablebolt#
Collardist.from
$
Drilledholelength
Cableboltlength
Cablebolttype
Installationmethod
Surfacefixturetype
Cableboltlength
Cablebolttype
Installationmethod
Surfacefixturetype
Groutbatch#
Groutflow
Comment#
3.9.1 Design Specifications
Specification: Cablebolt installation layout Example (See attached NOTES)
Plan #: 1450/-1 Stope #: 1400/C
Ring #: A
7 See attached Cablebolt drilling observation report.
Installation method:
A
....
Installation method, BT: Collar finishing, GF; W:C = 0.4; Additives, None.
As drilled Design specification Feedback: As installed
1 -2.8 11 10 PS A Pl
2 -2.3 11
3 -1.6 11
4 -0.8 11
5 0 11
6 0.9 10
7 1.9 11
...
...
A shaded box indicates an acceptable change from the original design.
Date cablebolts placed: Crew Leader:
Date cablebolts grouted: Crew Leader:
Date surface fixtures installed: Crew Leader:
See the attached Installation layout notes for additional information about this sheet.
In the feedback section of this sheet, please indicate where the installation followed
design with a check mark, 6, or record the change(s) made. Any comments about or
problems with the equipment or procedure, as well as feedback on the grouting
procedure must be noted on the Cablebolt installation observation report.
Design Specifications: 3.9.1 continued
Specification: Cablebolt installation layout NOTES
Collar dist.
from $
The hole collar distance from the $ is measured horizontally at the drift
back and along the sidewall.
Cablebolt
type
PS Plain 15.2 mm diameter strand.
TS Twin or double 15.2 mm diameter strand.
TS+Sp Twin 15.2 mm diameter strand with 56 mm diameter spacers
placed every 1 metre along the length of the cablebolts.
BC Single birdcaged strand.
TBC Twin birdcaged strand: two BC strands wired together.
DBC 14 wire birdcaged strand.
BA 25 mm diameter bulbed or nutcaged anchor formed from
15.2 mm diameter strand.
TBA Twin bulbed anchor: two BA strands offset and wired
together.
Installation
method
BT Uphole, breather tube installation.
UGT Uphole, grout tube installation
DGT Downhole, grout tube installation
Collar
finishing
F Fabric collar plug: burlap, cotton waste, shredded rag.
GF Grout soaked fabric collar plug: burlap, cotton waste, rag.
GP Grout collar plug.
WW Wooden wedge.
EF Expansive foam collar plug.
RP Rubber plug.
VP Victaulic pipe plug.
RSP Resin collar plug.
Surface
fixture type
Pl Plate of x by y surface dimensions and z thickness.
BPl Butterfly plate of x by y surface dimensions and z thickness.
DPl Domed plate of x by y surface dimensions and z thickness.
Str Strapping of x by y surface dimensions, z thickness, s spacing
between holes, and d hole diameter.
Grout
batch #
A record of the batch mixing details must be kept on the Cablebolt
installation observation report sheet. The grout batch # must be
recorded on the report sheet and the layout specification sheet.
Grout flow
observation
BT installations: Indicate the holes for which there was grout flow out of
the end of the breather tubes with a check mark, 6.
GT installations: Record the time between the appearance of watery
grout at the collar of the hole and of grout of design consistency, (min).
Indicate any holes for which grouting was incomplete with a cross, :.
Comment # Any comments about the performance of equipment, materials or
procedures should be noted on the Cablebolt installation observation
report sheet. Cross-reference to the layout sheet with a comment #.

Procedure: Cablebolt placement
1 CB-A: Prepare
the borehole.
A1 Upholes.
A2 Downholes.
2 CB-B: Prepare
the cablebolt
strand.
B1 Cablebolt strand in uncut coils on a palette.
B2 Pre-assembled, pre-cut cablebolts in coils on a
palette.
B3 Pre-assembled, pre-cut lengths of cablebolts.
B4 Cablebolt strand in uncut coils in a reel or dispenser to
be inserted directly into the borehole, prior to cutting
the strand.
B5 Cablebolt strand in uncut coils in a reel or dispenser to
be cut into lengths, prior to insertion in the hole.
3 CB-C:
Assemble the
cablebolt
element.
C1 Twin strand cablebolts with spacers.
C2 Twin strand cablebolts without spacers.
C3 Twin strand flared cablebolts.
4 CB-D: Attach
the hanger
D1 Spring steel hanger.
D2 Pre-attached nut for spring steel hanger.
D3 Bent wire hanger at the toe of the hole.
D4 Bent wire hanger at the collar of the hole.
D5 Pre-attached external fish hook hanger.
D6 No hanger required.
5 CB-E: Prepare
and attach the
tube(s).
E1 Grout tube to toe of hole; to be left in the hole during
and after grouting.
E2 Grout tube to toe of hole; to be retracted during
grouting.
E3 Breather tube to toe of hole, and short length of grout
tube in collar of hole.
6 CB-F: Place the
cablebolt
element.
F1 Upholes with breather tubes.
F2 Upholes or downholes with grout tubes; Grout tube to
be left in the hole during and after grouting.
be inserted with the cablebolt element, and to be
retracted from the borehole during grouting.
be inserted into each hole just prior to grouting, and to
be retracted from the borehole during grouting.
Procedure and Safety: 3.9.2 continued
Procedure: Cablebolt placement continued
7 CB-G: Finish the
borehole collar.
G1 Cotton waste or dry burlap collar plug.
G2 Grouted burlap collar plug.
G3 Grout collar plug.
G4 Wooden wedge.
G5 Rubber collar plug.
G6 Resin collar plug.
G7 Victaulic pipe collar plug.
G8 Expansive foam collar plug.
G9 No borehole collar finishing required.
Feedback:
- Any changes to the cablebolt length or type must be recorded on the Cablebolt
installation layout specification sheet.
- Any comments about the installation procedure, materials or equipment should
be recorded on the Cablebolt installation observation report sheet, and cross
referenced to the Cablebolt installation layout specification sheet.
Procedure: Grout mixing and pumping
1 CB-H: Mix the
grout.
H1 Paddle mixer, Drum mixer or Colloidal mixer.
2 CB-I: Clean the
grout mixer.
I1 Paddle mixer, Drum mixer or Colloidal mixer.
3 CB-J: Pump the
grout.
J1 Breather tube installation method.
J2 Grout tube installation method.
J3 Breather tube installation method in fractured
rockmass: fractured zone extent is known in advance.
J4 Grout tube installation method in fractured rockmass:
fractured zone expected during grout pumping.
J5 Grout tube installation method in fractured rockmass:
fractured zone extent is known in advance.
4 CB-K: Clean the
grout pump.
K1 Cleaning all grout pumps.
K2 Piston pump.
K3 Progressing cavity pump.

Procedure and Safety: 3.9.2 continued
Procedure: Grout mixing and pumping continued
Feedback:
- Any changes to the installation method must be recorded on the Cablebolt
installation layout sheet.
- The grout batch # and observations of the grout flow from the breather tube or
collar of the hole, as discussed on the Cablebolt installation layout Notes sheet,
must be recorded for each cablebolt on the Cablebolt installation layout sheet.
- Information about each grout batch must be recorded on the Installation
observation report sheet.
referenced to the Cablebolt installation layout sheets.
Procedure: Surface fixture installation
1 CB-L: Fixture
installation.
L1 Plain plates, Domed plates, Butterfly plates, Straps.
Feedback:
- Record any changes to the surface fixture type on the Installation layout sheet.
referenced to the Cablebolt installation layout sheet.
To create a cablebolt placement procedure for the specific conditions at the site,
determine which of the separate procedures listed in the preceding tables and in
the following text are required. Most of the procedures will likely require some
edits and additions to accurately reflect the conditions at the site, and text shown
in italics within the sample procedures should be checked to ensure that it is
accurate for the equipment and conditions at the site.
As an example, procedures A1, B1, D1, E3, F1 and G5 could be used to specify
the installation of a single plain strand cablebolt to be installed in an uphole.
Grouting procedures could be created from H1, I1, J1, K1 and K3.
The list of the cablebolt placement options given here may seem extensive, but
is most probably not even complete. The authors observed a great deal of
innovation with cablebolting procedures at the mine sites visited during the course
of the project. In a number of cases, a new technique had taken quite some time
and trial and error to develop, however the final product worked well enough to
justify the expenditure. Often development of the technique was necessitated by
failure or lack of efficiency of the old method.
CB-A: Borehole Preparation
A1: Upholes.
1) Make the working area safe by scaling down any loose rock, following safe
scaling practice.
2) Clear all debris from the floor of the working area.
3) Measure the length of the boreholes. Make a note of any holes that are
longer or shorter than the specified length on the Cablebolt installation
layout sheet. (If the specific hole length is critical for the design, specify a
hole length deviation that should not be exceeded. For example, if a 10 m
hole is 8 m in length, in a specific design this cablebolt might be useless,
and therefore not worth installing in the short hole. In this case the cablebolt
should be installed at a later time in a new hole drilled to the correct length.
To assess the hole length allowable deviation, consider the effect of the short
cablebolt on the performance of the cablebolt system, the likelihood of the
hole being redrilled, the effect of a shorter hole on the effectiveness of
modified geometry cablebolts, and any other factors relevant to your site.)
4) Make a note on the Cablebolt installation observation report sheet of any
boreholes which are producing water.
A2: Downholes.
1) Make the working area safe by scaling down any loose rock, following safe
scaling practice. Clear all debris from the floor of the working area.
2) Clean the floor of the drift around the cablebolt holes of any debris or dirt
that could fall into the borehole during the installation process.
3) Blow the boreholes clean and dry. If the holes are very dirty, fill them with
water and blow them dry again.
4) Measure the length of the boreholes. If a hole is greater than 1 metre too
long, the cablebolt could support the ore as well as the waste rock. In this
case, mark the hole as bad, make a note on the Cablebolt installation layout
sheet and do not install the cablebolt. (If a too short cablebolt will greatly
reduce the effectiveness of the cablebolt pattern, specify a hole length
deviation that should not be exceeded. To assess the allowable hole length
deviation that is specified here, consider the effect of the short cablebolt on
the performance of the cablebolt system, the likelihood of the hole being
redrilled, the effect of a shorter hole on the effectiveness of modified
geometry cablebolts, and any other factors relevant to the site.)

CB-B: Cablebolt Strand Preparation
B1: Cablebolt strand in uncut coils on a palette.
1) Place the coil of cablebolts in a clean, open working area away from other
people and equipment.
2) Stand in the middle of the coil, and cut the binding straps with an air
powered cutter in the correct order so that the coil of cablebolts will unroll
in a controlled manner. Ensure that no one is in the range of any springing
or lashing cablebolts.
3) Cut the length of cablebolts detailed on the Cablebolt installation layout
specification sheet, using an abrasive cutter. Wear a face shield and leather
gloves when cutting cablebolts, since small shards of metal and sparks are
created. Lay the cablebolts out on a clean, dry surface, such as a wooden
platform, a row of PVC tubes, or a series of saw horses. Where plates or
other surface fixtures are to be used on modified geometry cablebolt
elements, cut the strands so that the last 1 m at the collar end of the strand
is straight.
4) If the cablebolts are dirty or rusty, clean them with a wire brush or a
pressurized water jet. If it impossible to clean them, reject them.
B2: Pre-assembled, pre-cut cablebolts in coils on a palette.
1) Place the coil of cablebolts in a clean, open working area away from other
people and equipment.
2) Stand in the middle of the coil, and cut the binding straps with an air
powered cutter in the correct order so that the coil of cablebolts will unroll
in a controlled manner. Ensure that no one is in the range of any springing
or lashing cablebolts.
3) Separate apart the individual pre-assembled cablebolts, and lay them out on
a clean, dry surface, such as a wooden platform, a row of PVC tubes, or a
series of saw horses. Ensure that all components of the pre-assembled
element remain intact and are clean.
4) If any of the cablebolts are dirty or rusty, clean them with a wire brush or
a high pressure water jet. If it is not possible to clean them, reject them.
5) Where pre-assembled modified geometry elements are to have surface
fixtures attached after grouting, make sure that the end of the cablebolt
element with the straight section of strand will be used at the hole collar.
6) When hangers are included with the pre-assembled cablebolt elements, lay
the cablebolt out with the hanger closest to the drill hole collar.
CB-B: Cablebolt Strand Preparation continued
B3: Pre-assembled, pre-cut lengths of cablebolts.
1) Separate apart the individual pre-assembled cablebolts, and lay them out on
a clean, dry surface, such as a wooden platform, a row of PVC tubes, or a
series of saw horses. Ensure that all components of the pre-assembled
element remain intact and clean.
2) Where pre-assembled flared cablebolt elements are to have surface fixtures
attached after grouting, ensure that the end of the cablebolt element with the
straight section(s) of strand is furthest away from the borehole collar.
3) When hangers are included with the pre-assembled cablebolt elements, lay
the cablebolts out with the hanger end nearest the collar of the borehole.
4) If the cablebolts are dirty or rusty, clean them with a wire brush or a high
pressure water jet. If it is not possible to clean the cablebolts, reject them.
B4: Cablebolt strand in uncut coils in a reel or dispenser to be inserted directly
into the borehole, prior to cutting.
1) Place the reel or dispenser in a clean area as near to the working face as
possible. Ensure that the cablebolt strand will not drag on the floor while
being dispensed.
2) If the strand within the coil is rusty, check with the engineer to see whether
the surface can be cleaned, or whether the whole coil should be rejected.
3) After the hanger and tubes have been attached to the cablebolt, and the
cablebolt has been inserted into the hole (F), cut off the end of the steel
strand using an abrasive cutter. Wear a face shield and leather gloves.
B5: Cablebolt strand in uncut coils in a reel or dispenser to be cut into lengths,
prior to insertion in the hole.
1) Place the reel or dispenser in a clean area as near to the working area as
possible. Ensure that the strand will not drag on the floor while being
dispensed.
2) If the strand within the coil is rusty, check with the engineer to see whether
the surface can be cleaned, or whether the whole coil should be rejected.
3) Pull the required length of cablebolt from the dispenser onto a clean, dry
platform, and cut using a tungsten grinder blade. Wear a face shield and
leather gloves when cutting cablebolts, since small shards of metal and
sparks are created. If plates or other surface fixtures are to be used on
modified geometry cablebolt strand, cut the strands so that the last 1 m of
the strand at the collar end of the strand is straight.

CB-C: Cablebolt Element Assembly
C1: Twin strand cablebolts with spacers.
1) Place two individual lengths of strand
on a clean, dry working space.
2) Insert the cablebolt strands into the 56
mm diameter green plastic double
spacers. The first spacer should be
placed at 0.5 m from the toe end of the
cablebolts. The rest of the spacers
should be placed every 1 m along the
length of the cablebolts. The last spacer
should be placed 0.5 m from the collar
position.
C2: Twin strand cablebolts without spacers.
1) Place two individual lengths of strand on a clean, dry working space.
2) For twin strands to be installed in upholes, place the cablebolts so that the
ends of the separate strands are offset slightly. The protruding strand should
be long enough to accept the hanger.
3) Wire the toe end of the cablebolts together in two places.
C3: Twin strand flared cablebolts.
1) Place two individual lengths of strand
on a clean, dry working space.
2) Position the individual strands so that
the flared sections are offset at even
spacings along the length of the
cablebolt element.
3) Wire the toe end of the cablebolts
together, leaving enough space for the
hanger. Tie the strands together with
wire every 2 metres.
CB-D: Attachment of Hanger
D1: Spring steel hanger.
1) Wear leather gloves as the sharp steel
strips can cut your hands.
2) Ensure that the spring steel strips are
evenly spaced around the outside of the
hanger, and that the nuts are tight.
3) Attach the spring steel hanger to the
end of the cablebolt using a steel band
hose clamp, so that the spring steel
strips on the top of the hanger are above
the end of the cablebolt.
D2: Pre-attached bolt for spring steel hanger.
1) Wear leather gloves as the sharp steel
strips can cut your hands.
2) Remove the nut from the bolt.
3) Slide 3 7.5 cm by 2 cm steel strips over
the end of the bolt, spacing the strips
evenly around the outside of the hanger.
4) Place the washer over the strips.
5) Screw the nut back onto the end of the
bolt and tighten.
D3: Bent wire hanger at the toe of the hole.
1) Working at the end of the cablebolt
nearest the hole collar, bend over a 75
mm length of one wire so that it forms
a hook, using the bending tool. The
hook should point towards the collar
end of the cablebolt and be at a 40- to
50- angle away from the cablebolt
strand.
2) The hook must be long enough to grip
the borehole wall.

CB-D: Attachment of Hanger continued
D4: Bent wire hanger at the collar of the hole. (This type of hanger is not
applicable when surface fixtures are to be used, and may not be very
effective in certain circumstances.)
1) Working at the end of the cablebolt
furthest from the hole collar, bend 75
mm of one wire of the end of the
cablebolt away from the strand, using
the bending tool. The hook should point
towards the collar end of the cablebolt
and be between 40- and 50- away from
the cablebolt element.
2) The hook must be long enough to grip
the borehole wall when it is inserted.
D5: Pre-attached external fish hook hanger.
1) Using the bending tool, bend out the
wires of the pre-attached hanger. The
bent wire should be between 40- and
50- from the cablebolt.
2) The wire(s) must be long enough to
grip the borehole wall. If the wire(s)
are too long, they may not be at a great
enough angle from the cablebolt to grip
the hole well.
D6: No hanger required.
The hanger must be strong enough to support the full weight of the ungrouted
cablebolt, and be long enough to grip the borehole wall securely. If the borehole
diameter has changed from design, or if the bent wires are shorter or longer than
specified, the hanger may fail to support the cablebolt, leading to very dangerous
working conditions. The selection of adequate hangers is discussed in greater
detail in Section 2.11.4.
CB-E: Tube Preparation and Attachment
E1: Grout tube to toe of hole: to be left in the hole during and after grouting.
1) Cut the end of the 19 mm I.D. grout
tube at a 45 angle, using a hack saw.o
2) Cut two 10 cm long slots into the side
of the grout tube. This is especially
important for downhole cablebolts
where the end of the tube can become
easily plugged with cuttings.
3) Position the grout tube beneath the arm
of the hanger for protection. Using
duct tape, attach the grout tube to the
cablebolt element, so that the end of the
grout tube is 15 cm to 30 cm from the
toe end of the cablebolt.
4) Straighten the tubes so there are no
kinks or twists in the plastic.
E2: Grout tube to toe of hole: to be retracted during grouting.
1) Cut the end of the 19 mm I.D. grout tube at a 45 angle, using a hack saw.o
E3: Breather tube to toe of hole, and short length of grout tube at the collar
of the hole.
1) Cut the end of the 10 mm I.D. breather
tube and the 19 mm I.D. grout tube at a
45 angle, using a hack saw.o
2) Position the breather tube beneath the
arm of the hanger for protection. Using
duct tape, attach the breather tube to the
cablebolt element, so that the end of the
breather tube is 15 cm below the toe
end of the bolt.
3) Mark the collar position on the
cablebolt. When surface fixtures are to
be used, 1 metre of one strand should
be below this mark.
4) Straighten the tubes so that there are no
kinks or twists in the plastic.

CB-F: Cablebolt Placement
F1: Upholes with breather tubes.
1) Remove any obstructions or loose near
the borehole collar.
2) Insert the assembled cablebolt element
with attached breather tube into the
hole. Adopt safe lifting practices
when pushing the cablebolts into the
hole, and don't overexert yourself. One
crew member must guide the free end
of the cablebolt. While pushing the
cablebolt into the hole, ensure that the
breather tube is not crushed, kinked or
twisted. Stop pushing the cablebolt
when the marked collar position is still
1 m away from the rock face.
3) If it is impossible to insert the cablebolt
into the hole, push a grout tube up the
hole and flush the hole with water. If
the hole is still obstructed, redrill it, and
then place the cablebolt.
4) Insert the grout tube a few cms into the
collar of the hole. Tape the grout tube
to the cablebolt element, leaving the breather tube free.
5) Pull on the cablebolt to check that it is securely anchored.
6) Flush water through the breather tube to clean the borehole and cablebolt.
7) Place a plastic cap on the end of the cablebolt.
8A) If cotton waste or burlap is to be used to seal the collar of the hole, leave the
cablebolt protruding from the hole, and proceed to CB-G: the Borehole
collar finishing instructions.
8B) If cement or expansive foam collar packing are to be used to seal the collar,
push the cablebolt the rest of the way into the hole.
9) Cut off the protruding end of the breather tube, leaving enough length of
tube outside the hole to allow easy observation of the return of the grout,
and to tie off the tube after grouting is finished.
10) Cut off the protruding end of the grout tube, leaving enough length of tube
outside the hole to tie off the tube after grouting is finished.
11) Roll up the tubing that is left protruding from the hole so that it is out of the
way until you start grouting.
CB-F: Cablebolt Placement continued
F2: Upholes with grout tubes: Grout tube to be left in the borehole during and
after grouting.
1) Remove any obstructions or loose rock
from the borehole collar.
2) Insert the assembled cablebolt element
with grout tube into the hole. Adopt
safe lifting practices when pushing
the cablebolt into the hole. Someone
must guide the end of the cablebolt, so
that it does not whip around in the air.
While pushing the cablebolt into the
hole, ensure that the grout tube is not
crushed or kinked. Try to keep the
grout tube straight, as twists and bends
in the tube will create points where
grout pressure builds up.
3) Cut off the protruding end of the grout
tube, leaving enough length of tube
outside the hole to be attached to the
grout pump hose, and to tie off the tube
when grouting is finished. Roll the end
of the tube up and secure it out of the
way.
4) Pull on the cablebolt to ensure that it is securely anchored.
5) Place a plastic cap on the end of the cablebolt. Then if the cablebolt does
fall from the hole, there will be less chance of damage.
F3: Upholes or downholes with grout tubes: Grout tube to be inserted with the
cablebolt element and to be retracted from the borehole during grouting.
1) Remove any loose or obstructions from around the borehole collar.
2) Holding the end of the grout tube at the end of the cablebolt element, insert
both into the hole. Adopt safe lifting practices when pushing the cablebolts
into the hole. Ensure that the reel of grout tubing is able to unroll easily,
and the tubing will not kink or be crushed during the cablebolt placement.
Keep the grout tube as straight as possible and prevent it from twisting as it
enters the hole. Roll up the rest of the tube and secure it out of the way.
3) Pull on the cablebolt to make sure that it is securely anchored.

CB-F: Cablebolt Placement continued
F4: Upholes or downholes with grout tubes: Grout tube to be inserted in each
hole just prior to grouting and to be retracted from the borehole during
grouting.
1) Remove any loose or obstructions from around the borehole collar.
2) Insert the cablebolt element into the hole. Adopt safe lifting practices
when pushing the cablebolts into the hole. Someone must guide the end of
the cablebolt, so that it does not whip around in the air.
3) Pull on the cablebolt to check that it is securely anchored.
4) The grout tube will be inserted into the hole during the grouting procedure.
However, after the first couple of cablebolts have been placed in a working
area, attempt to insert the grout tube into the boreholes. If there are
problems with inserting the grout tube into the hole, stop installing the
cablebolts in that area, and inform the underground supervisor or engineer
about the problem.
As is mentioned in all of these cablebolt placement procedures, it is very
important that the grout and/or breather tubes are not damaged during insertion of
the element into the borehole. Damage is most likely to occur when thin walled,
weak tubing is used. It is worth spending a few extra cents for each unit length of
tubing to reduce the chances of damaging the tube. Damaged tubing may not
transmit the grout during pumping, thereby potentially rendering the cablebolt
installation in that borehole useless.
Some mine cablebolting crews observed by the authors had devised reels for
dispensing the tubing, for much increased ease of handling. At these sites, one of
the crew members inserted the steel strand into the borehole, while another
member guided the tubing into the hole, keeping it straight and free of twists and
kinks.
CB-G: Borehole Collar Finishing
G1: Cotton waste or dry burlap collar plug.
1) Wrap cotton waste or burlap around and between the cablebolt, grout tube
and breather tube, above the collar position of the cablebolt element. 20 cm
of the cablebolt above the collar position mark should be fabric covered.
2) Push the cablebolt the rest of the way into the hole, tamping the material
into the borehole collar tightly, while keeping the cablebolt(s) in the centre
of the hole. Double cablebolts that will both be plated or strapped, must be
far enough apart so that a barrel can fit over each of them.
tube outside the hole to allow easy observation of the return of the grout
down the tube, and to tie off the tube when grouting is finished.
4) Cut off the protruding end of the grout tube, leaving enough tube outside the
hole to be attached to the grout hose, and to tie off the tube after grouting.
5) Blow on the end of the grout tube. You should feel air coming out of the
end of the breather tube. If you do not feel air return, push some more
fabric into the collar of the hole. If this doesn't lead to air return, the rock
around the collar or along the borehole could be fractured. Do not grout the
hole, and inform the engineer at the end of the shift.
6) Before grouting, flush water through the breather tube to saturate the fabric.
G2: Grouted burlap collar plug.
1) Dip the burlap into a bucket of thin cement grout paste. Wear safety glasses,
long sleeves and water proof gloves to prevent cement grout burns.
2) Wrap the burlap around and between the cablebolt, grout tube and breather
tube at and above the collar position of the cablebolt element. At least a 20
cm length of the cablebolt above the marked collar position should be
covered by the burlap. Double cablebolts that will both be plated or
strapped, must be far enough apart so that a barrel can fit over each of them.
3) Push the cablebolt the rest of the way into the hole, tamping the burlap into
the borehole collar tightly, keeping the cablebolt centred in the borehole.
tube outside the hole to allow easy observation of the return of the grout
down the tube, and to tie off the tube when grouting is finished.
5) Cut off the protruding end of the grout tube, leaving enough tube outside the
hole to be attached to the grout hose, and to tie off the tube after grouting.
6) Blow on the end of the grout tube. You should feel air coming out of the
breather tube. If you do not feel air return, push some more material into
the collar of the hole. If you still don't get air return, the rock around the
borehole could be fractured. Do not grout the hole, and inform the engineer.

CB-G: Borehole Collar Finishing continued
G3: Grout collar plug.
1) Insert a short section of grout tube into the collar of the hole. This section
must be shorter than the grout tube that was inserted with the cablebolt
element. Mark this grout tube, so that there will be no confusion between
the two tubes. Make sure that the distance between the end of the longer
grout tube and the collar plug is sufficient to create a large enough grout
plug to support the full weight of the grout column.
2) Plug a short section of the collar of the hole with cotton waste or burlap,
making sure that the cablebolt remains in the centre of the borehole.
3) Mix 0.40 W:C grout following the procedures given in CB-H.
4) Pump the grout into the hole through the shortest length of grout tube,
following the procedures given in CB-J. Continue to pump until the grout
flows back down the longer grout tube. Stop pumping the grout and kink
over and tie off the shortest grout tube. Wash off the end of the longer
length of grout tube, so that there will be no problems with attaching the
grout pump hose to the tube when you return to grout the borehole.
5) Blow air through the longer grout tube to ensure that it is still open. If it is
plugged, pump a small amount of water into the tube to flush it.
6) Ensure that the grout collar plug is placed early enough so that it has cured
to an adequate strength to support the grout column when the cablebolt hole
is grouted.
G4: Wooden wedge.
1) Push a wooden wedge into the collar of
the hole to secure the cablebolt until
ready to grout. The wedge should be
placed between the cablebolt and the
borehole wall, with the tubes pushed
aside.
2) Make sure that the grout and/or breather
tube(s) are not crushed or pushed out of
round during this operation.
3) Pull on the end of the cablebolt to ensure
that the wedge is secure.
4) Take a great deal of care with this operation, especially if no hanger has
been used at the toe of the hole to provide added support to the cablebolt
element.
G5: Rubber collar plug (used at Mines Gaspé and Brunswick Mine, Canada).
Note: This method must be used with a work platform that can be raised,
such as a scissor lift.
1) Place a wooden wedge in the borehole
collar to secure the cablebolt (procedure
CB-G4). Remove the wedge when
ready for grouting.
2) Remove any obstructions or loose rock
from the collar area, making the collar
rock as smooth as possible.
3) Place the rubber plug onto the end of
the 3/4 diameter pipe or rod that will
be used to support it. Hold the end of
the pipe or rod vertically with the lower
end resting on the scissor lift platform.
4) Raise the platform until the rubber plug
is tightly wedged into the hole collar.
During this process, feed the breather
tube into the slot on the side of the
rubber plug.
5) If the rock around the collar of the
borehole is rough or jagged, use some
waste cotton around the rubber plug to help seal the collar.
G6: Resin collar plug (used at Macassa Mine, Canada).
1) Wear rubber gloves, long sleeved clothing and safety glasses to prevent resin
burns. Never stand under the hole.
2) Lay a piece of paper towel or flexible rag out on the work space. Cut open
a tube of resin and spread it out on the paper towel, mixing the two
components of the resin well.
3) Wrap the resin soaked paper or fabric around the collar position of the
cablebolt, and push the cablebolt the rest of the way into the borehole.
4) Push more paper towel or rag into the collar to completely seal the hole.
The resin sets very quickly, so that it is possible to grout the hole almost
immediately after sealing the collar.

G7: Victaulic pipe collar plug (used at Campbell Mine, Canada: Bourchier et
al, 1992). The cablebolt is inserted into the hole through the pipe collar.
1) Remove any loose or obstructions near
the collar of the borehole.
2) Attach the 2 Victaulic clamp to the
grooved end of the pipe casing.
3) Cut a 2 slot on the plain end of the
pipe casing with a hack saw.
4) Crimp the slotted end of the pipe so that
it fits into the hole.
5) Insert the crimped end of the pipe into
the borehole and tap with a hammer
until the pipe bead is inserted into the
hole. Insert the cablebolt into the hole.
6) Remove the Victaulic clamp.
7) Coat the inside of the Victaulic reducer
with a thin layer of grease.
8) Push the breather tube through the hole
on the inside of the reducer.
9) Slide the reducer along the breather
tube, and fasten to the casing using the
2 Victaulic clamp, which has been
coated with grease.
10) Tighten the clamp.
11) Coat the outside of the clamp and
reducer with grease.
12) 24 hours after grouting the hole, the victaulic clamp and reducer can be
removed from the pipe and used again.
G8: Expansive foam collar plug.
1) The foam collar plug should be sprayed into the hole following the
manufacturer's instructions. Ensure that the length of the plug is adequate
to hold the grout column.
2) Wait at least 24 hours before starting to grout.
G9: No borehole collar finishing required.
CB-H: Grout Mixing
H1: Paddle, Drum and Colloidal mixers.
1) Determine how much water will be needed for the mix dependent upon the
size of the cement bags and the grout mix design.
Size of cement
bag (kg)
Volume of water required (litres)
W:C = 0.3 W:C = 0.35 W:C = 0.4 W:C = 0.45
10 3 3.5 4 4.5
25 7.5 8.75 10 11.25
40 12 14 16 18
2) Grout can burn exposed skin severely. Wear glasses, gloves and protective
clothing when working with cement powder and grout. Barrier cream
should be used on any exposed skin, and on your hands. If any grout does
touch bare skin, the area should be washed with soap and water and
completely rinsed. If grout gets into your eye, flush with fresh water and
report to the first aid station immediately.
3) Wear a dust mask when breaking open bags of cement.
4) Cement should be stored on water proofed, plastic wrapped palettes close to
the intended pump set up. You will need about 80 kg of cement for each 15
m borehole.
5) Ensure that the mixer and pump are clean with no dried grout lumps.
6) Provide detailed instructions for assembling the mixer components and
connecting the system to the air supply (see manufacturers instructions).
7) Use whip-checks on all air hose connections, and blow the hoses clean
before connecting.
8) Check to ensure all valves are off.
9) Turn the air on at the header.
10) Fill the hopper with water. Turn on the mixer to check that it is operating.
Flush water through the grout pump. Remove all water from both.
11) Measure and pour the required amount of water for the batch. It is very
important to mix the specified water:cement ratio. If the mix is too dry, it
will be hard to pump and may lead to pressure build up in the grout tubes
and eventually bursting of the grout tubes. If the mix is too wet, the grout
strength with be reduced and the grout will be more fluid and will more
easily flow out of the hole collar or into rockmass fractures.
12) Turn on the mixer.
13) Where specified, mix the correct volume of the wet additive into the water.
Agitate the water and additive slowly to prevent foaming.

CB-H: Grout Mixing continued
14) If the mixer and pump are a single unit, start to recirculate the water through
the pump and back into the mixing hopper. Recirculating the grout will
assist in complete mixing, and lead to a smoother mix. If the pump doesn't
have a recirculating valve, place the outlet end of the grout pump hose into
the mixer hopper, keeping it free of the mixer blades, and start the pump.
15) Break each bag of cement powder on a screen placed on top of the hopper
of the mixer. Add the dry powder to the water slowly, and make sure that
any lumps of pre-hydrated cement are removed. If there are a lot of lumps,
discard the bag, and note the problem on the Cablebolt installation
observation report sheet.
16) Where specified, add the correct volume of dry additive.
17) Mix the grout completely so that there are no lumps, and the grout has a
smooth consistency. There should be no lumps visible on the surface of the
mix. Stop the mixer briefly and remove any build up of dry cement or
lumps on the paddles or bin walls.
18) Whenever the mixer will be idle for more than 1/2 hour (lunch time, shift
end), clean it following the instructions given in procedure CB-I.
CB-I: Grout Mixer clean up
I1: Paddle, Drum and Colloidal grout mixers.
Mixer clean up is essential for all types of mixers to ensure efficient and clean
operation. Detailed instructions for the clean up of the mixer should have been
provided by the supplier when the mixer was purchased. Add these instructions
to this procedure. The time after which the grout will become too difficult to clean
will depend on the type of cement and on the additives in use.
Work with the crew to ensure that the instructions given in the procedure result
in a completely clean mixer. In general though, it is important to:
1) Remove as much of the grout from the mixer as possible.
2) Fill the mixer hopper with fresh water and operate the mixer.
3) Pour this water out of the hopper.
4) Wash the hopper and paddles with more fresh water.
5) Scrape any accumulations of grout from all exposed surfaces.
6) After cleaning the mixer, fill it with fresh water and leave it full of water
until the next time that it is used.
CB-J: Grout Pumping
J1: Breather Tube Installation Method.
touch bare skin, wash the area with soap and water and rinse completely.
If grout gets into your eye, flush with fresh water, and then report to the first
aid station immediately.
2) Set up the grout pump as close as possible to the boreholes to minimize the
length of grout tube required. Longer tubes result in greater fluid pressure
in the tubes and higher power requirements for the pump.
3) Add instructions for the regular maintenance of the pump. These details
might include such items as lubrication or air filter replacement.
4A) If you have not been recirculating the grout during the mixing process,
pump grout from the grout pump hose onto the floor until it is the same
consistency as the grout in the hopper. The first portion of the grout may be
watery and must be emptied from the hose before grouting the hole.
4B) If you have been recirculating the grout through the grout pump hose back
into the hopper, shut off the grout pump.
4C) If you have been recirculating the grout through a recirculation valve during
mixing, turn off the valve, slow the grout pump down, but don't stop it.
5) Clean any grout off the outside of the grout pump hose, and attach the hose
to the grout tube using the quick screw connector or the modified vice grips.
6) Turn the pump on slowly. The pump speed must be fast enough to force the
grout through the tube, but not so fast that excess pressure builds up and
bursts the plastic tubing, and not so slow that the pump stalls.
7) Place the end of the breather tube in a bucket of water. Air will bubble
through the water as it is forced out of the hole by the grout front.
8) Estimate the volume of grout required to fill the hole from Figure 3.9.1.
9) Monitor the level of the grout in the pump hopper. If the grout volume that
has been pumped into the hole is more than twice what it should be, the
rockmass is fractured. Stop grouting, remove the collar plug and disconnect
the grout hose from the grout tube to allow the grout to drain from the hole.
If you are unable to remove the collar plug, pump water through the grout
tube until you see clean water flowing from the end of the breather tube.
Wait for at least 24 hours before returning to plug the collar again and grout
the hole. Record the problem on the Cablebolt installation observation
report, and make a note of the grout volume initially pumped into the hole.
10) Try to block any grout leaks out of boreholes or cracks with foam or cotton
waste. Alert the engineer at the end of the shift about the problem.
11) Keep pumping until grout of design consistency flows from the end of the
breather tube. Stop the grout pump.

Figure 3.9.1: Volume of grout required per metre of cablebolt.
CB-J: Grout Pumping continued
12) Kink over the end of the breather and grout tubes and tie them off.
13) Disconnect the grout tube from the grout pump hose.
14) Start the grout recirculating in the pump if you are not immediately ready
to start grouting the next borehole.
15) Use all of each batch before starting to mix more grout. Never mix grout
continuously. Do not start to grout a hole if there is not enough grout to fill
it. Discard the remaining grout and mix a fresh batch.
16) If there are any delays which prevent the grout from being pumped into the
boreholes within 1/2 hour of the start of the mixing time, pump the grout out
onto the floor of the working area away from any downholes, and clean the
grout pump (Procedure CB-K). If additives or high early strength grout are
being used, the time that the grout can remain in the pump will be less.
17) Whenever the pump will be idle for more than 1/2 hour, make sure that the
pump is completely clean and is left full of clean water.
18) Never stand under newly grouted boreholes. Any water dripping out of the
holes can burn skin badly.
19) Where surface fixtures are to be used, wait 24 hours and then follow the
instructions in Procedure CB-L. If surface fixtures are not to be used, trim
the ends of all cablebolts that could create a safety hazard.
20) 24 hours or more after grouting, cut off the breather and grout tubes.
J2: Grout Tube Installation Method.
touch bare skin, the area should be washed with soap and water and
completely rinsed. If grout gets into your eye, flush with fresh water, and
then report to the first aid station immediately.
2) Set up the grout pump as close as possible to the boreholes to minimize the
length of grout tube required. Longer tubes result in greater fluid pressure
in the tubes and higher power requirements for the pump
3) Estimate the volume of grout required for the borehole from Figure 3.9.1.
4A) If you have not been recirculating the grout during the mixing process,
pump grout from the grout pump hose until it is the same consistency as the
grout in the hopper. The first portion of the grout may be watery and must
be emptied from the hose before grouting of the hole starts.
4B) If you have been recirculating the grout through the grout pump hose back
into the hopper, shut off the grout pump.
4C) If you have been recirculating the grout through a recirculation valve during
mixing, shut off the valve, and slow the grout pump down, but don't stop it.
5) Clean any grout off the outside of the grout pump hose, and attach the hose
to the grout tube using the quick connector or modified vice grips.
6) Start pumping the grout slowly. The pump speed must be fast enough to
force the grout through the tube, without stalling, but not so fast that excess
pressure builds up and bursts the plastic tubing.
7) Monitor the level of grout in the hopper. It is very important that a
progressing cavity pump is never operated dry for more than 15 seconds. If
the grout level stops dropping in the hopper while pumping, stop the pump.
Remove the grout hose from the grout tube and try to restore normal
operation. If the grout has become too stiff to flow into the pump, add water
to the hopper and pump the waste cement out onto the floor. If the pump
stalls, and it is impossible to finish grouting the hole, estimate how much
grout has already been pumped into the borehole and make a note. Stop the
pump and detach the grout pump hose from the grout tube. Try to pump
grout through the hose. If the hose is jammed, stop pumping the grout, and
try to determine where the blockage has occurred. Manipulate the grout
pump hose in the area of the blockage to try to free it. If all else fails,
discard the hose and start with a new hose. If grout flows freely from the
pump hose, it is likely that the grout froze in the hole due to the presence
of fractured rock. Follow the instructions for grouting in fractured rock
given in Procedure CB-J4.

8) Keep pumping until a donut of grout of design consistency squeezes out
from the collar of the borehole. A thin cement slurry will precede the
thicker grout front. Do no stop pumping until the thicker grout
appears. Determine the time between the first appearance of grout, which
may be watery, and the appearance of design consistency grout. Record this
time on the Cablebolt installation layout sheet. If a tongue of grout which
appears solid but does not completely encircle the cable (i.e. a donut),
keep pumping until the exiting grout forms a complete encapsulating ring.
Make a note in the installation report of any unusual grout flow.
9) If grout has appeared at the borehole collar before the expected full volume
of grout has been pumped into the hole, make a note of how much grout has
been pumped into the borehole on the Cablebolt installation report sheet,
and inform the engineer of the problem.
10) Stop the grout pump.
11) Kink over the end of the grout tube and tie it off.
12) Disconnect the grout tube from the grout pump hose.
13) Start the grout recirculating again in the pump if you are not immediately
ready to start grouting the next borehole.
14) Use all of each batch before starting to mix more grout. Never mix grout
continuously.
15) Monitor the volume of grout in the hopper. If it appears that you will not
have enough grout to fill the next borehole, discard the remaining grout and
mix a fresh batch.
16) If there are any delays which prevent all of the grout from being pumped
into the boreholes within 1/2 hour of the start of the mixing time, pump the
grout out onto the floor of the working area, and clean the grout pump using
the procedures given in Procedure CB-K. If there are additives in the grout
mix, or if high early strength grout is being used, the time that the grout can
remain in the pump will be less.
17) Whenever the pump will be idle for more than 1/2 hour, make sure that the
pump is completely clean and is left full of clean water.
18) Never stand under newly grouted boreholes. Any water dripping out of the
holes can burn skin badly.
19) Where surface fixtures are to be used, wait 24 hours and then follow the
instructions in Procedure CB-L. If surface fixtures are not to be used, trim
any cablebolts protruding from the holes that could create a safety hazard.
20) 24 hours or more after the grouting has been completed, cut off the kinked
over grout tubes.
J3: Breather tube installation method in fractured rockmass: fractured zone
extent known in advance.
Follow normal grouting procedures (CB-J1), with these additional instructions.
This procedure will work when the longer grout tube has been inserted past the
fractured section of the borehole. If grout continues to be lost into the rockmass
when pumping through the longer grout tube, use a longer grout tube (#2) or even
a third (longer still) as required.
1) When installing the cablebolt, insert
two grout tubes into the borehole. The
first grout tube, #1, should be inserted a
short distance into the collar of the
borehole following normal procedure.
The second grout tube, #2, should
extend just beyond the expected
position of the fractured rockmass zone.
2) From Figure 3.9.1, estimate how much
grout is required to fill the borehole
between the collar and the end of the
longer grout tube, #2, and between the
end of grout tube #2 and the toe.
3) Pump grout through grout tube #1.
Monitor the grout level in the hopper.
4) Stop grouting as soon as good quality
grout flows out of the longer grout tube,
#2, or after you have pumped twice as
much grout into the hole as should be
required to fill the hole.
5) Tie off grout tube #1.
6) Drain the grout out of grout tube #2.
Clean and blow air in the end of the
tube. If it is blocked, pump a small
amount of water to clear the tube.
7) Detach grout tube #1 from the grout
pump hose, and continue on to grout the next borehole or clean the pump.
8) Return after 24 hours, and pump grout through grout tube #2.
9) Keep pumping until grout of design consistency returns along the breather
tube. Stop the grout pump, and follow normal procedures as in CB-J1.
10) If grout is flowing through the fractures and obstructing the grouting of
other boreholes, drill, install and grout each cablebolt individually.

J4: Grout tube installation method in fractured rockmass: fractured zone
encountered during grout pumping.
The thicker grouts used in the grout tube method should flow along the
borehole more readily than into the open discontinuities in the rockmass.
Therefore there should be little problem with grouting in fractured rock with the
grout tube method and grouts of W:C 0.35. However there have been reports
of the thicker grout freezing in uphole installations in very fractured rock
(Oliver, 1992) due to water loss. This happens when the water lubricating the
flow of the grout along the borehole wall flows away into the rockmass leaving
the grout too dry to flow well. There are two options in this case. The first option
is to push a second piece of grout tube into the borehole until it reaches the
position of the frozen grout front, and to continue to grout the hole with the new
tube. The second option involves washing the grout out of the hole and the grout
tube, and trying to regrout the hole normally. In this case, it is essential that the
grout is washed from the hole by water introduced at the collar, and that water is
not pumped into the grout tube until the hole is basically clear of grout.
J5: Grout tube installation method in fractured rockmass: fractured zone
recognized in advance of pumping.
In locations where the grout front frequently freezes in the borehole, the
conductivity of the open discontinuities must be reduced. This can be achieved
by spraying a thin grout mixture onto the borehole walls with a slotted pipe
attached to the end of the grout pump hose. Manufacture a short length of pipe
with a sealed upper end, and thin slots cut at numerous positions around the
circumference of the pipe. Attach a connector end to the lower end of the pipe to
attach a grout tube. Pump a relatively thin grout slurry out through this pipe. Once
the grout slurry layer has set, normal grout pumping procedures should work
(Oliver, P., personal communication, 1993).
CB-K: Grout Pump Clean up
K1: Cleaning all grout pumps
A complete cleanup of the grout pump is essential to ensure efficient and clean
operation, and to reduce the maintenance required. Detailed instructions for the
clean up of the grout pump should have been provided by the pump supplier. Add
these instructions to this procedure. Work with the crew to ensure that the
instructions given in the procedure result in a completely clean pump.
CB-K: Grout Pump Clean up continued
K2: Cleaning the Piston pump
Whenever the pump will be idle for more than 15 minutes:
1) Pump as much of the grout from the hopper as possible.
2) Before the pump runs completely dry, fill the hopper with fresh water.
3) Continue filling the hopper with water and operating the pump until the
water flowing from the end of the grout hose is clear.
4) Scrape any accumulations of grout from all exposed surfaces of the pump.
At the end of each shift:
5) Remove the foot valve by unscrewing with a pipe wrench. Wash the ball
and ball stop.
6) Remove the shaft casing or riser tube taking care not to damage or bend it.
Wash the inside of the tube.
7) Re-assemble the pump and pump fresh water through it.
8) Blow air through the grout hose.
K3: Cleaning the Progressing cavity pump
Whenever the pump will be idle for more than 15 minutes:
1) Pump out all unused grout, but do not operate the pump dry.
2) Wash down the hopper, paddles and housing. Add water to the hopper and
continue to pump until the water flowing from the grout hose is clear.
3) Leave the hopper full of water until ready to pump the next batch.
At the end of the shift:
4) Disconnect the grout hose, open the end clamps and remove the rotor/stator
and auger. Disassemble the rotor/stator and wash all components.
5) Blow air through the grout hose.
6) Reassemble the pump, and fill the hopper with water.
CB-L: Surface Fixture Installation
L1: Plain plates, Domed plates, Butterfly plates, Straps.
1) Collect the wedges and barrels. If there are different types of wedges and
barrels in use on site, ensure that the wedges match the barrels. When the
wedges are placed inside the barrel without the cablebolt, the wedges should
fit snugly and be of the same taper angle as the barrel, should not stick out
from the bottom of the barrel, and should be flush with or slightly
protruding from the top.

CB-L: Surface Fixture Installation continued
2) Do not use any rusty or oily wedges or barrels. If there is mud on the
wedges or barrels, or the wedge teeth contain grit, clean them completely.
3) Collect the plates and/or straps. Ensure that the plates are of the correct
type, surface dimensions, thickness and shape. The straps must have holes
of the specified diameter at the correct spacing and be of the specified
thickness and surface dimensions.
4) Make sure that the hydraulic jack used to install the wedges and barrels is
ready. If the jack has a spring load nose cone, check it every so often. It
should be almost impossible to compress the spring with your hand. If the
spring is soft, replace it. Clean any rust or dirt from the jaws of the jack.
5) Clean the end of the cablebolt, so that there is no mud, dried cement grout
or other substance on the wires. Cut off any exposed tubing.
6A) If using single strand, place the strap or plate over the end of the cablebolt.
6B) For double strand cablebolts, bend over or cut off the end of one of the
cablebolt strands. Place the strap or plate over the end of the other cablebolt.
7) Fit the strap over all of the cablebolts that it will be used on to check that it
will fit. It is important to do this before installing any of the wedges and
barrels so that any problems will be identified before you start. If some of
the cablebolts will not fit through the pre-drilled holes in the strap, move
onto the next group of cablebolts and inform the engineer of the problem at
the end of the shift.
8) Ensure that the plate or strap is roughly perpendicular to the cablebolt. If
the cablebolt is angled more than 25- away from perpendicular to the rock
surface, use a spherical washer between the plate and the barrel, or use a
rounded barrel and domed plate to allow the cablebolt to remain straight.
Note where spherical elements have been used on the Cablebolt installation
layout sheet. Kinking the cablebolt sharply ( 25-) at the collar of the hole
will damage, weaken or break the wire(s).
9) Make sure that the dome on the domed plates will be facing away from the
rock face, so that the flat part of the plate rests on the rock and the dome can
be compressed.
10) Position the barrel over the cablebolt, against the plate or strap.
11) Place the wedges over the end of the cablebolt and slide up to the barrel.
Make sure that the individual wedges are evenly spaced, and not touching.
12) Using the jack, tension the cablebolt to the load specified, making sure that
no part of the jack is bearing against the wedges, with the exception of the
setting spring. The jack should only bear against the barrel, otherwise the
wedges may split or fracture. Take care that the jack is always supported,
especially when working on overhead cablebolts. Add any instructions for
the operation of the particular jack.
13) If any of the plates or straps look at all loose, inform the engineer.
Installation Trouble Shooting
Problems will arise during the cablebolt installation process. Suggestions for
solutions to some of the problems that might arise during installation are given
here. A number of suggestions are provided by Pakalnis et al. (1991).
1) Borehole blocked so that it is impossible to insert the cablebolt.
Try to dislodge the obstruction with a grout tube. If it is possible to insert the
grout tube to the end of the hole, flush the hole with water. Redrill the hole if it
is impossible to clear the obstruction. If hole obstruction is a frequent problem, try
to insert and grout the cablebolt as soon as possible after drilling is finished.
2) Grout pump won't run.
Check for any restricted air lines, blockages in the grout hose, closed valves or
dirty filters.
Check the air supply pressure and volume.
3) Borehole collar leaks during grouting.
Check to ensure that the grout is of the correct consistency. If it is too thin,
work with the crew until the correct grout consistency is being mixed and pumped.
Alter the collar plugging method for breather tube installations if the plug
continues to leak.
4) No air flowing from breather tube, and grout pump stalling at the start of hole
grouting, indicating that the breather tube is blocked.
Ensure that the breather tube is placed at least 15 cm below the toe of the hole.
Flush the breather tube with water prior to grouting to remove any blockages.
Use a stronger tube to prevent pinching of the tube or the collapse of the tube
under pressure (Cluett, 1991). Increase the diameter of the breather tube.
5) No air flowing from the breather tube, and grout pumping well, indicating that
the rockmass around the borehole is fractured.
Stop grouting, drain the grout from the tubes and hole, and allow the grout to
set in the fractures. Return in 24 hours or later to finish grouting the hole.

3.9.3 Quality Control
Quality control guideline: Cablebolt placement
Quality Control Consequences of poor
quality control
control
Good Poor
Clean
cablebolt
Dirty
cablebolt
The cablebolt / grout bond
will be reduced by mud, oil
or dirt coating the wires of
the cablebolt strand.
Store the cablebolts in a
clean, dry area until ready for
use. Keep the cablebolt
cutting / assembling area
clean and dry. Support the
cablebolts off the ground, or
dispense them from a reel or
carousel.
Cable
hanger
well
attached
and/or of
the
correct
diameter
Cable
hanger
poorly
attached
and/or of
the
wrong
diameter
The cablebolt could fall from
the hole before the grouting
is complete, creating a safety
hazard.
Ensure that the hanger is
strong enough to hold the
cablebolt in the hole, is
securely attached and is of
the correct diameter for the
borehole. Bent wire hangers
should be of the correct
length, and bent out to the
correct angle.
Breather
or grout
tube near
the end
of hole
Breather
or grout
tube not
to end of
the hole
If the breather or grout tube
is too short in uphole
installations, then the toe of
the hole will be left empty of
grout. If either tube is right at
the end of the hole, grout
flow will be difficult.
Attach the breather or grout
tube at least 15 cm or 6
inches from the toe end of
the cablebolt.
Cablebolt
centred
in the
hole
Cablebolt
at the
edge of
the hole
The cablebolt circumference
will not be completely
embedded in the grout, likely
resulting in reduced capacity.
Use spacers every 1 m along
the length of plain strand
cablebolts and attach them
securely. The cages of
modified geometry
cablebolts keep the
cablebolt away from the
edge of the hole.
Hole
collar
fully
plugged
Hole
collar not
well
plugged
Grout will leak out of the
hole, leaving voids in the
grout column. The leaking
grout can burn anyone
working below the hole
collar. A well designed and
installed plug may leak if
very wet grout is installed.
Establish a collar plugging
method that is adequate to
completely seal the collar
and can withstand the
pressure exerted by the
grout column on the plug.
The method must also be
quick and easy to
accomplish.
Quality Control: 3.9.3 continued
Quality control guideline: Grout mixing
quality control
control
Good Poor
Dry,
fresh
cement
Lumpy
cement
The overall cement grout
strength will be reduced due
to partial early hydration.
The grout mixer and pump
can be damaged by lumps in
the grout as well. Lumps
lodged in the grout or
breather tube or within the
pump will prevent complete
grouting of the hole.
Ensure that cement bags are
well water proofed during
transport and storage, by
wrapping the palettes in
plastic. Store the bags of
cement in a dry, shaded,
clean place until required for
use. Remove any lumps
from the cement powder
before placing it in the mixer,
and discard any bags with a
lot of lumps. Do not stack
palettes more than 2 high.
Grout of
correct
W:C
Grout too
wet
Cablebolt capacity will be
reduced from design level.
Corrosion resistance of the
grout is reduced. Grout will
flow out of the borehole in
the uphole grout tube
installation method, or leak
past the collar plug in the
breather tube method.
Ensure that the specified
quantities of grout and water
are placed in the mixer. Any
mixing problems should not
be rectified by adding more
water to the mix, but should
be brought to the attention of
the underground supervisor
and the engineer, so that the
grouting system or grout mix
design can be modified.Grout too
dry
Cablebolt capacity may be
reduced from design level, if
air voids are entrained in the
grout column when the grout
is too dry to flow properly.
The mixing and pumping
systems may not function
adequately.
Grout
with
correct
volume
of
additive
Too
much
additive
The grout strength may be
reduced if too much additive
is added, and/or the grout
may not flow as designed.
Ensure that the additive is
easy to measure out, such
as by using a scoop for dry
powder or a measuring
container for wet liquid. A
possible solution to problems
with controlling additive
volume is to purchase the
grout with dry additive added
to the mix at the plant before
bagging.
Not
enough
additive
The grout will not flow as
designed, and may hang up
in the borehole during
pumping or may leave voids
in the grout column.

Quality control guideline: Grout pumping
quality control
control
Good Poor
Grout
fully
mixed
Grout not
fully
mixed
If the grout is not completely
mixed, it will be harder to
pump the grout mix, and the
grout strength will vary along
the length of the hole.
Ensure that the grout is
completely mixed, with no
lumps and dry patches. Mix
in a batch and not
continuously, to ensure
consistent water:cement
ratio and grout strength. If
the grout was not
recirculated during mixing,
pump the grout through the
pump hose onto the floor
until a consistent, non-watery
grout is flowing, before
starting to grout the hole.
Breather
or grout
tube near
the end
of the
hole
Breather
or grout
tube at
the end
of the
hole
If the breather or grout tube
is pushed right to the end of
the hole, grout flow will be
impeded. The grout will not
flow down the breather tube,
or the grout tube might burst.
Attach the breather or grout
tube at least 15 cm or 6
inches from the toe end of
the cablebolt, during the
placement of the cablebolt.
Cut the end of the tubes at a
45 angle.o
Breather
/ grout
tubes
fully
grouted
Breather
/ grout
tubes not
fully
grouted
If the tubes are not fully filled
with grout, then a void will be
left in the column of grout.
The larger this void, the
lower the cablebolt load
carrying capacity
Ensure that the grout returns
along the breather tube.
When grouting is complete,
bend over the ends of the
breather and/or grout tubes,
and tie securely.
Borehole
filled with
grout
Borehole
not full
of grout
Any gaps in the grout column
will reduce the capacity of
the cablebolt system. These
gaps could be caused by
grout slumping or blobbing
down the hole in uphole toe
to collar installations, by
grout leakage at the collar of
any uphole installation, grout
leakage into a fractured
rockmass in any hole
orientation, or by segregation
of the grout column (for
grouts of water:cement ratio
greater than 0.4).
Match the grout flowability to
the installation system and
the cablebolt geometry to try
to ensure that the
impedances to full column
grouting are minimal (Figure
2.5.9. This can be checked
with a pipe pumping test.
Make sure that it is possible
to mix and pump the grout
water:cement ratio specified
in design and that the correct
W:C is being installed.
Quality control guideline: Surface fixture installation
quality control
control
Good Poor
Cablebolt
and
fixtures
clean
Cablebolt
and
fixtures
dirty or
rusty
The strength of a rusty cable
or rusty fixtures will likely be
weakened, and mud or rust
will reduce the frictional
interaction between the
wedges and cable.
Use clean, fresh cablebolts
and fixtures.
Free end
of
cablebolt
long
enough:
inside
and
outside
the hole
Free end
of
cablebolt
too short:
inside
and
outside
the hole
Inside the hole: If there is no
free length of cablebolt
inside the hole, only a small
% of the tensioning load
applied will remain.
Debond a section of the
cablebolt near the collar of
the hole.
Outside the hole: A short
piece of cablebolt outside
the hole will not leave
anything for the tensioning
jack to grip on to.
Ensure that a long section of
cablebolt is protruding from
the hole after insertion. Do
not cut the cablebolt off until
after the surface fixture has
been installed.
Cablebolt
free
length
perpen-
dicular
to rock
face
Non 90-
angle
between
cablebolt
and plate
Geometric mismatch
between the cablebolt and
the rock surface will create a
sharp bend in the cablebolt
element during tensioning.
Some of the strands in the
cablebolt may be broken.
Design the cablebolt pattern
so that the hole angle is
roughly perpendicular to the
rock surface. The maximum
angular mismatch should be
no more than 25-.
Barrel
and
wedges
fit
together
well
Barrel
and
wedges
mis-
matched
Mismatched wedges and
barrels will not fit well
together, and may not retain
the load applied during
plating, or the steel wedges
or barrel may chip or split
due to eccentric loading.
Ensure that matched
wedges and barrels remain
together. If there are several
different sizes and types in
use, make sure that each
kind is very clearly and
distinctly marked.
Correct
plate
Wrong
plate
If the plate is too small, thin,
or soft, then the face
restraint on the rockmass
may not be adequate.
Use the plates designed for
the job, and ensure that the
plates purchased are stiff
and strong enough.
Tension
jack
good
Tension
jack too
soft
If the spring in the nose cone
is soft, the tension applied by
the jack will be lost.
Check the spring stiffness in
the nose cone of the jack
periodically. Replace soft
springs immediately. The
teeth in the grips should also
be sharp and clean.

3.9.4 Feedback
Feedback: Cablebolt installation observation report
Plan #: Stope #:
Cablebolt ring #s:
Date: Crew Leader:
Grout
batch
#
W:C Grout
additive
Batch
mixing
time
Any problems with cement quality
or mixing equipment?
Hole #'s grouted from this batch.
Com-
ment #
Ring
#
Hole
#
Batch
#
Deviation from design, problems with
equipment, materials or installation procedure?
Changes to the layout should be commented upon on this sheet, and noted on the
Cablebolt installation layout sheet.
3.10 Automated Cablebolting Systems
The design specifications, installation procedures and quality control guidelines
discussed in Sections 3.7 through 3.9 are generally applicable to the installation
of cablebolts with automated systems. However, it will be necessary to modify
the sheets to reflect the specific requirements and procedures for automated
equipment. Some suggestions for modifications to the sheets are given below for
automated cablebolt installation equipment.
The automated cablebolting machines grout the borehole first, and then insert
the cablebolt into the grout filled hole. It is essential that no part of the operation
results in substantial voids being introduced into the grout column. For example,
if the cablebolt is secured in the hole by a hanger created by pulling a section of
the cablebolt out of the hole, kinking the cablebolt, and then pushing it back into
the hole, grout will be lost from the hole (Bawden, Hyett and Cortolezzis, 1992).
In addition, as was mentioned in Chapter 2, the grouting equipment supplied with
the automatic cablebolting machine must be capable of completely mixing a batch
of W:C 0.35 grout and of pumping it into the longest borehole that will be used
in the cablebolt patterns at the mine site.
The main difference between the automated systems and the regular
cablebolting procedure is that one person per shift operates the equipment and is
responsible for the entire cablebolting job, from drilling the hole to inserting the
cablebolt into the grout filled borehole. The operator works under a protective
canopy at all times. The operator has the benefit of drilling the boreholes and
installing the cablebolts, thereby gaining continuous experience with the rockmass
in the working area. It is still very important that the operator record the drilling
and installation feedback information as suggested in Sections 3.8.4 and 3.9.4, so
that the engineer can monitor changes to the cablebolt design and rockmass
conditions. It is likely that a different crew will install the surface fixtures where
they are used.
The manufacturers of automated systems should provide operational instruction
in the form of manuals and hands on training sessions. All of the relevant,
machine-specific instructions should be added to the sheets in this section.
3.10.1 Automated System Design Specifications
The cablebolt design specifications given to the cablebolting machine operator,
should include modified versions of the Cablebolt Layout Plan and Section
(Section 3.8.1) and Cablebolt Installation Layout and Notes (Section 3.9.1).
The Cablebolt layout plan and section specification sheets will be directly
applicable to most uses of automated cablebolting equipment. However a number

Cablebolt#
Holeanglefrom
$
Collardist.from
$
Drilledholelength
Cableboltlength
Surfacefixturetype
Cableboltlength
Surfacefixturetype
Groutbatch#
Groutflownotes
Comment#
of changes to the Cablebolt Installation Layout and Notes will have to be made.
For example, there is usually no variation in the cablebolt type specified in the
design, since individual, specialized pusher heads are required for different types
of modified cablebolts.
Specification: Automated cablebolt installation layout (See attached NOTES)
Plan #: Stope #:
Cablebolt ring #:
See attached Cablebolt drilling observation sheet.
Cablebolt type:
As drilled Design
specification
Feedback: As installed
1
2
3
4
5
6
7
8
9
10
....
Date cablebolts grouted and placed: Operator:
Date surface fixtures installed: Crew Leader:
See attached Installation layout NOTES for additional information about this sheet.
In the feedback section of this sheet, please indicate where the installation followed
design with a check mark, T, or record any change(s) made. Any comments about
or problems with the equipment or procedure, as well as feedback on the grouting
procedure must be noted on the Cablebolt installation observation report.
Automated System Design Specifications:
3.10.1 continued
Specification: Automated system cablebolt installation layout NOTES
Collar dist.
or angle
from $
The hole collar distance is measured horizontally from the $.
The position is specified as angle from vertical $. (+ CW; - CCW)
Cablebolt
type
PS Plain 15.2 mm diameter strand.
TS Twin 15.2 mm diameter strand.
NC 25 mm diameter nutcaged 15.2 mm diameter strand.
BA 25 mm diameter bulbed anchor formed from 15.2 mm
diameter strand.
Surface
fixture type
Pl Plate of x by y surface dimensions and z thickness.
BPl Butterfly plate of x by y surface dimensions and z thickness.
DPl Domed plate of x by y surface dimensions and z thickness.
Str Strapping of x by y surface dimensions, z thickness and s
spacing between holes of diameter d.
Grout
batch #
A record of the batch mixing details must be kept on the Cablebolt
installation observation report sheet. The grout batch # must be
recorded on the report sheet and the layout specification sheet.
Grout flow
observation
Record the time between the appearance of watery grout at the collar of
the hole and of grout of design consistency, (sec).
Note if there is any flow of grout from adjacent boreholes during
grouting, or from the borehole at any time after grouting is finished or
after the cablebolt has been inserted into the hole.
Comment # Any comments about the performance of equipment, materials or
procedures should be noted on the Cablebolt installation observation
report sheet. Cross-reference to the layout sheet with a comment #.
3.10.2 Automated System Procedure and Safety
The cablebolt installation and grouting procedures must include any safety
guidelines applicable to the operation of the automated system. The order of work
presented in the installation procedures provided in this book for manual
installations will have to be altered for the automated equipment, because the
operator first drills the borehole, then grouts the borehole and finally inserts the
cablebolt(s) into the grout filled hole. Some samples of modified procedure sheets
are given in the following text. It is essential that any routine maintenance
procedures, as well as commonly encountered problems and their solutions be
added.

Procedure: Automated system grout mixing and pumping and
cablebolt placement
1 Spray the boom with oil before starting the grouting operation. This will help to
prevent the grout from sticking to the machine.
2 If you are working in very fractured ground, run the cablebolt up the hole first, to
clear away any loose rock that could obstruct the hole. If the rockmass is very
fractured, drill the boreholes for one ring of cablebolts, then install the grout and
cablebolts immediately in that ring.
3 Place the required amount of water into the mixer hopper.
4 Start the mixer blades rotating, and slowly add the required number of cement
bags. Keep mixing until the grout is of a smooth consistency. Make sure that the
quantities of water and cement in the mix are correct. The mixer blades must be
kept rotating at all times while the batch is in the hopper.
5 Pump grout through the grout tube until you see design consistency grout flowing
from the end of the tube.
6 Stop the grout pump, and insert the grout tube to the toe of the borehole.
7 Turn on the high pressure water jet on the boom to clean the outside surface of
the grout tube as it is being withdrawn from the hole.
8 Start the grout pump and automatic withdrawal of the grout tube. If you see the
grout tube starting to buckle, it is not being retracted quickly enough.
9 When the grout begins to flow from the collar of the borehole, the grout tube
should be about 0.5 to 1 metre above the collar of the hole. Once the grout tube
has been withdrawn completely from the hole, stop the pump. If the grout
continues to flow from the borehole after pumping has stopped, the grout is too
wet, either due to poor quality control on the mix or due to the presence of
flowing water in the borehole. Note the problem on the cablebolt installation
observation report. If the problem is severe, the grouted hole cannot be used.
10 Within 15 minutes of any grouting system break down, wash the pump and flush
the tube with fresh water.
11 Insert the cablebolt into the grout filled hole. Keep the head of the boom within
about 1 foot of the face, to minimize the free length of cablebolt, thereby
reducing the possibility of kinking the cablebolt during insertion.
12 Periodically check the cablebolt reel under the machine. If the cablebolt looks
rusty, you will probably encounter difficulties with the flow of the cablebolt off the
reel. If you run into problems with the cablebolt feed, change the reel. Place the
bound cablebolt pack into the cassette and then cut the steel binding straps.
13 When the cablebolt is about 3 metres from the toe of the hole, use the ram to
kink the cablebolt. It is important that the cablebolt is not retracted too much
from the hole during this operation, since this will result in grout loss. (Note that
the process of securing the cablebolt in the hole can be quite different from site
to site, depending upon the specific operation of each machine).
14 Stop pushing the cablebolt once it has reached the toe of the hole.
15 Cut the cablebolt off using the cutter head.
Procedure: Automated system clean up
1 Wash the mixer thoroughly using the high pressure water jet. Flush the water
and left over cement out through the bypass valve at the bottom of the mixer.
2 Fill the pump hopper with fresh water and flush water through the grout tube
system until clear water is running from the end of the tube. Leave the hopper
full of clean water after all of the tubes have been flushed clean.
3 Wash the boom and grout tube reel with the high pressure water jet to remove
all cement.
3.10.3 Automated System Quality Control
Many of the quality control guidelines listed for conventional cablebolt
installations are applicable to an automated installation. Modify the following
sheets as required:
Borehole preparation quality control guidelines 3.8.3
Cablebolt placement quality control guidelines 3.9.3
Grout mixing quality control guidelines 3.9.3
Grout pumping quality control guidelines 3.9.3
Surface fixture installation quality control guidelines 3.9.3
3.10.4 Automated System Feedback
The installation feedback is equally important for conventional and automated
cablebolt installation procedures. Modify the following sheets for your site and
usage of the cablebolting machine:
Drilling observation report 3.8.4
Cablebolt installation layout and notes 3.9.1
Cablebolt installation observation report 3.9.4
A number of sites using automated cablebolters were visited by the authors. In
all cases, an extensive period of adaptation, machine customization and procedural
adjustment was necessary before adequate efficiency was achieved.

3.11 Quality Control Monitoring and Testing
The maintenance of good procedures and quality during the cablebolt
installation is critical and of great importance to safety and economics. Poor
quality control can compromise the load carrying capacity of the cablebolts
substantially.
The crews should be monitoring the quality of the installation and reporting any
problems that they encounter (Section 3.5). The quality of the installation must
also be monitored by the underground supervisors, engineers and technicians
during spot check visits to the work area while the crew are installing the
cablebolts. Methods for monitoring the quality during installation are given in
Section 3.11.2. In addition, the quality of the installations should be checked after
the installation is complete, following the guidelines given in Section 3.11.3.
If quality control problems are encountered, steps must be taken immediately
to rectify the problems (Section 3.12). If the cablebolt system's performance is
strongly compromised by the quality control problems, then additional cablebolts
should be installed in the area.
3.11.1 Effect of Quality Control on Cablebolt Capacity
A cablebolt must be well installed in order to attain the load carrying capacity
used in design: the clean, undamaged cablebolt strand must be completely
surrounded by a full grout column of design water:cement ratio.
The load carrying capacity of cablebolts can be greatly reduced by poor quality
control during installation. The capacity reductions that have been quantified in
laboratory pull tests on cablebolts are listed in Table 3.11.1. Other factors which
have not been quantified, but which can reduce the capacity of a cablebolt are
listed in Table 3.11.2. Decreasing stress in the rockmass around the cablebolts can
severely reduce their load carrying capacity as well (Chapter 2).
Table 3.11.1: Cablebolt capacity reductions due to poor installation quality control.
Muddy, greasy or heavily rusted cablebolts (Lappalainen and
Pulkkinen, 1982; Leclair, 1995)
-50%to -70%
Ungrouted breather tube (11 mm I.D.: W:C = 0.45) (Goris, 1990) -30%
Excess water in the grout mixture: (Reichert et al, 1992)
W:C increased from 0.35 to 0.40
W:C increased from 0.40 to 0.55
-15%
-45%
Grout column incomplete Up to -100%
If poor quality control during installation has been reported in an area, and is
expected to reduce the cablebolt system capacity to an unacceptably low level,
additional cablebolts should be installed.
Table 3.11.2: Impact of quality control problems
Reduction of cablebolt system capacity due to poor quality control
1)reduction of the interaction between cablebolts in the pattern,
2)reduction of the capacity of an individual cablebolt, or
3) reduction of the grout quality or the bond strength.
Borehole Incorrect collar position or borehole orientation 1
Incorrect borehole length or diameter 2
Dirty or wet borehole 3
Cablebolt strand Incorrect type or capacity 2
Incorrect length 2
Dirty, very rusty or oily 3
Cablebolt
element
placement
Cablebolt not to end of borehole 2
Cablebolt not central in the borehole 2
Grout tube or breather tube not at specified length 2
Grout mixing
and pumping
Incorrect cement type 3
Incorrect water:cement ratio 3
Inadequate mixing time or mixing continuously 3
Breather or grout tubes not full of grout 2
Hole not filled with grout or grout segregation 2
Surface fixture
installation
Incorrect or mismatched wedge and barrel fittings 2
Incorrect or mismatched surface hardware 2
Acute angle of the cablebolt with the face 2
Incorrect installed tension 2
Plate not in full contact with the rock surface 2
Subsequent
effects
Blast damage 3
Corrosion 2
Machine damage 2
Stress decrease 2

3.11.2 Checking Quality Control during Installation
Quality control can be checked during the cablebolt installation by:
1) Observing the installation process.
2) Inspecting the grout quality.
3) Testing the grout water:cement ratio.
4) Testing the grout strength.
Each of these checks should be made frequently, during visits to the working
areas. The results of the checks should be discussed with the crews: either to
reinforce that they are doing a good job, or to solve any problems that are
identified.
Observation of Installation Practice
The drilling and cablebolt crew members are required to check certain aspects
of the installation quality control while they are working, and to report on any
problems with the installation procedure, materials or equipment.
The underground supervisor, the rock mechanics technician and the engineer
should visit the work site often, to monitor the quality of the cablebolt installation,
and to identify any problems with the installation procedure. These observations
can be recorded on the Cablebolt quality control check list, which follows.
Modify the sample check list given here to represent the procedures and equipment
in use at your site. Any specific information, shown in italics on the sample sheet,
about the borehole diameter, the grout mix design or any other item should be
added to the sheet. It would be a good exercise to have all mine personnel
involved with the cablebolting procedure complete this form from time to time to
ensure that they remain aware of the potential quality control problems.
Any problems with the installation procedure that are observed should be
discussed with the crew immediately, and should also be noted in the stope file.
When a new crew is formed, or new equipment or hardware is implemented,
it is important to check the quality of the installation more frequently than usual.
Even when the quality control seems to be excellent, and there are apparently no
problems, it is important to continue to check the installation quality control every
so often.
Cablebolt Quality Control Check List:
Breather Tube Installations
Page 1 of 2
Your Name: Date:
Crew Location:
Please 7for a Yes answer to the following questions. Deviations from design practice or additional
comments should be noted in the boxes on the right side of this form.
Drilling Comments
Is the specified drill bit being used: Diameter = 65 mm?
Are the holes all drilled to the design length: 10 m?
Are the holes dry? If not, what quantity of water is flowing (l/s)?
Material storage, transport and handling
Are the cablebolts clean? If not, please specify the problem:
± - muddy, oily or greasy?
± - slightly rusty? (rough to the touch)
± - very rusty? (slippery to the touch)
Are the cement bags free of lumps?
Are the cablebolts free of kinks and damage?
Are the breather or grout tubes kinked or damaged?
Cablebolt placement
Are dirty, oily or rusty cablebolts cleaned before placement?
Are the boreholes clean and dry before cable placement?
Are two plain 7 strand cablebolts being placed in every hole?
Is the spring steel hanger attached securely?
Are 56 mm spacers being placed every 1 m along the cablebolts?
Are 13 mm ID breather and 19 mm ID grout tubes being used?
Is the breather tube attached .15 mm below the cablebolt toe end?
Is there at least 1 m of grout tube inside the hole?
Is each cablebolt 11 m long, with 1 m sticking out of the hole?
Are all of the flared sections of the cablebolt inside the hole?
Is the cablebolt secure in the hole?
Is the collar of the hole being well sealed with cotton waste?

Cablebolt Quality Control Check List: Breather tube Page 2 of 2
Your Name: Date:
Crew Location:
Grout mixing Comments
Consistent, reliable mixer operation? If not, does the mixer:
± - sound laboured?
± - ever stall?
How many 25 kg cement bags being used in the mix?
Are 40 litres of water being used in the mix?
Is the grout being mixed in individual batches?
Is the grout mixture smooth, well mixed and free of lumps?
Is the grout being mixed for at least 15 minutes before pumping?
Is the grout mixer completely clean and left full of water after use?
Grout pumping
Is the grout pump hose / grout tube connection adequate?
Are the grout tubes free from leaks or bursting?
Does good quality grout always return down the breather tubes?
Does one grout batch fill 2 holes?
Are the collar plugs free of leaks?
Are the breather tubes kinked and tied off as soon as grouting stops?
Are the grout tubes kinked and tied off as soon as grouting stops?
Is the grout pump completely clean and left full of water after use?
Plating and strapping
Are the cablebolts roughly perpendicular to the rock surface?
Are the wedges and barrels free of rust, dirt and oil?
Do the wedges and barrels fit well together?
Are the plates tight after installation of the wedge and barrels?
Any other comments or problems observed, or suggestions for improvement?
Cablebolt Quality Control Check List:
Grout Tube Installations
Page 1 of 2
Your Name: Date:
Crew Location:
Please 7for a Yes answer to the following questions. Deviations from design practice or additional
comments should be noted in the boxes on the right side of this form.
Drilling Comments
Is the specified drill bit being used: Diameter = 65 mm?
Are the holes all drilled to the design length of 10 m?
Are the downhole collars covered to prevent water inflow?
Are the upholes dry? If not, what quantity of water is flowing (l/s)?
Material storage, transport and handling
Are the cablebolts clean? If not, please specify the problem:
± - muddy, oily or greasy?
± - slightly rusty? (rough to the touch)
± - very rusty? (slippery to the touch)
Are the cement bags free of lumps?
Are the cablebolts free of kinks and damage?
Are the breather or grout tubes kinked or damaged?
Cablebolt placement
Are dirty, oily or rusty cablebolts cleaned before placement?
Are the boreholes clean and dry before cable placement?
Is a single plain 7 strand cablebolt being placed in every hole?
Is a spring steel centralizer used on the end of each cablebolt?
Is a 19 mm ID grout tube being used?
Is each grout tube attached .15 mm from the cablebolt toe end?
Is each cablebolt 11 m long, with 1 m sticking out of the hole?
Are wooden wedges placed between the cablebolt and hole collar?
Is each grout tube round and free of distortion by the wedge?
Are the cablebolts secure in the hole?
Are at least 2 m of grout tube left hanging from each hole collar?

Cablebolt Quality Control Check List: Grout tube Page 2 of 2
Your Name: Date:
Crew Location:
Grout mixing Comments
Consistent, reliable mixer operation? If not, does the mixer:
± - sound laboured?
± - ever stall?
How many 25 kg cement bags being used in the mix?
Are 35 litres of water being used in the mix?
Is the grout being mixed in individual batches?
Is the grout mixture smooth, well mixed and free of lumps?
Is the grout being mixed for at least 20 minutes before pumping?
Is the grout mixer completely clean and left full of water after use?
Grout pumping
Is the grout pump hose / grout tube connection adequate?
Are the grout tubes free from leaks or bursting?
Does watery grout flow from the holes at first?
Is pumping continued until thick grout flows from the collar?
Does one grout batch fill 2 holes?
Are the borehole collars free of grout leaks after grouting stops?
Are the grout tubes kinked and tied off as soon as grouting stops?
Is the grout pump completely clean and left full of water after use?
Plating and strapping
Are the cablebolts roughly perpendicular to the rock surface?
Are the wedges and barrels free of rust, dirt and oil?
Do the wedges and barrels fit well together?
Are the plates tight after installation of the wedge and barrels?
Any other comments or problems observed, or suggestions for improvement?
Visual Inspection of Grout Quality
A visual inspection of the grout mixture will give a fairly accurate estimate of
the water:cement ratio (W:C) of the grout. The appearance of the grout at
different water:cement values can be estimated from descriptions and photos. It
is a good idea to prepare photos of grout with the crew and any other people who
will be checking the quality control, so that they can see grout of different
water:cement ratios. It is essential to make your own set of photos for any grout
mixes that include additives, since the visual appearance of grout mixtures can be
quite different. Add any specific details to the description of the grout mixtures
given in Table 3.11.3 that would be useful.
Always wear long water proof gloves when handling grout mixtures.
Table 3.11.3 Grout characteristics
Grout characteristics (after Hyett et al, 1992)
water:cement
ratio
Grout characteristics at end of
grout hose
Grout characteristics when
handled
0.30 Dry, stiff sausage structure. Sausage fractures when bent,
and grout is too dry to stick to
your hand. Grout can be rolled
into balls.
0.30 Moist sausage structure, which
melts slightly with time.
Sausage is fully flexible and
grout will stick to your hand.
Grout is easily rolled into wet,
soft balls.
0.35 Wet sausage structure, which
melts away with time.
Grout sticks readily to your hand,
even when upturned.
0.40 Sausage structure is lost
immediately. Grout flows
viscously under its own weight to
form a pancake.
Grout sticks readily to your hand,
but can be shaken free.
0.50 Grout flows readily and splashes
on impact with the ground.
Grout will drip from hand, with no
shaking required.

Figure 3.11.1: Visual characteristics of grout - an aid to quality control
Visual Characteristics of Grout
W:C = 0.35
W:C = 0.40
W:C = 0.45
W:C = 0.50
{ }
{ }
[ ] [ ]( ) { }
[ ] { }
{ }
Assume that water has a specific gravity, , of 1.0
(anything else and you shouldn' t be using it for grouting!):
Specific gravity of cement, . . 3.15 (Hyett et al., 1992)
mass of grout in kg
volume of grout in litres
mass of cement in kg
mass of water in kg
water:cement ratio
S G
S G
M M M
V V V
A
S G M kg V litres
S G
B M kg A
W C
B
A
W
C
grt grt cont cont
grt cont w
C grt grt
C
grt
. .
. .
. .
:
( ) ( ) ( )
( ) ( ) ( )
( ) ( )
( )
≈
= −
= −
=
−
−
= −
=
+
1
Grout Water:Cement Ratio Testing
The water:cement ratio of the grout can be easily measured by taking samples
of the grout during visits to the underground working location, using the following
procedure (Rheault, J., 1993, personal communication). Be aware, however, of the
inherent scatter in wet cement paste density (Figure 2.5.2).
1) Purchase some 1 litre plastic containers with screw top lids. Weigh the
containers without their lids. This value is M .(cont)
2) Completely fill each plastic container with water measured from a graduated
cylinder to determine the volume of the container = V .(cont)
3) Fill the container with grout taken from the end of the grout hose, mid way
through pumping a batch of grout. Close the lid of the container firmly.
Always wear long water proof gloves when handling grout mixtures.
4) Weigh the grout filled container without the lid. This value is M .(grt+cont)
5) Fill any voids in the container with a volume of water measured from a
graduated cylinder. The volume of the extra water added is V .(w)
6) Calculate the water:cement ratio, W:C, of the grout, by solving the
following equations:

Figure 3.11.2: Results of UCS testing of field grout samples from mine sites (after
Gendron et al, 1992).
Grout Strength Testing
The strength of the grout used in the cablebolting holes can be measured by
testing the strength of samples collected at the working face, using the following
procedure (Gendron et al, 1992). Noranda uses custom built, foam rubber lined
aluminum sample cases for storing and transporting the grout samples. The grout
samples are collected in PVC tubes (2 inside diameter and 8 length) with plastic
screw threaded end caps. Six of these tubes fit tightly into the case.
1) Spray the inside of the PVC tube with silicone spray, so that the grout sample
can be extracted easily from the tube prior to testing.
2) Take random samples of the grout mid way through pumping a batch. Always
wear long water proof gloves when handling grout mixtures.
3) Transport the samples to the surface. Try not to disturb the samples during
transport, and certainly do not turn the sample tubes upside down.
4) Send the cylinders for testing at any time after 7 days of undisturbed hydration.
Indicate the date each sample was taken, and when it should be tested.
5) Plot all results on a graph, such as Figure 3.11.1, and note the location of any
cablebolts grouted with low strength grout. If the grout is very weak
throughout an area, more cablebolts should be installed.
6) Inform the crew of the results. If the tests consistently indicate weak grout,
work with the crew to determine the source of the problem.
3.11.3 Checking Quality Control after Installation
Post-Installation Visual Inspection of Cablebolts
It is more difficult to determine if there have been installation quality control
problems during a post-installation inspection of the work area than from
observations made during the installation. However, Table 3.11.4 suggests some
of the items that can be checked after installation has been completed.
Table 3.11.4: Visual post-installation quality control checks.
Item to be checked Method for checking
Grout quality Look at any grout splashed onto the floor or walls of the work
area. If the grout looks more fluid than usual, the
water:cement ratio was too high.
Completeness of
grout column
Check the end of breather and/or grout tubes to see if they are
full of grout. If possible, cut off the end of grout or breather
tubes when making this check. If the tubes are not full of
grout, then you should spend some time observing the crew as
they install the cablebolts, since they may not be getting return
on the breather tubes, or may be using grout that is too thin
with the grout tube installation method.
Check any visible borehole collars for completeness of
grouting.
Plates and straps Check the angle of the cablebolt with respect to the rock
surface. The cablebolt behind the surface fixture should be
roughly perpendicular to the surface. If it is not perpendicular,
check the individual wires in the cablebolt strand for damage.
Observe the plates and straps. They should be of the correct
size and should be flush with the rock surface. You should not
be able to move the fixtures with your hand. If the fixtures are
loose, then they can be shaken off by blast vibrations, or will
be too loose to perform their intended function.
Check the orientation of any non-flat plates (e.g. domed or
butterfly plates). The rounded part of the plates should be
facing towards you (convex) and not in towards the rock face
(concave).
Wedges and
barrels
Try to move the barrel. It should be impossible to move it.
Observe the wedges and barrels. The wedges should be
protruding slightly past the end of the barrel, and should show
no signs of damage. There should be no grout, oil or dirt on
any of the wedge or barrel surfaces.

Post-Installation Cablebolt Pull Testing
Short lengths of cablebolts can be pull tested after installation. In these tests,
a load is applied to the cablebolt and the resulting deformation is monitored. A
detailed description of pull test procedures is given in Section 2.2.2. Pull tests for
comparison of cablebolt bond strength to published test results should be carried
out on cablebolts which are installed with 250 mm or 300 mm of grout bond. Such
tests can provide site specific capacity estimates. Where stress change and in
particular, stress relaxation is suspected, it is useful to perform field pull tests at
different phases of the mining sequence (before and after stope excavation, for
example). This will assess the risk of bond capacity reduction due to stress
decrease (Section 2.6.2).
Pull testing of cablebolts bonded over their full length will only detect major
deficiencies in the installation of a cablebolt. If a cablebolt is well installed with
good quality grout, for example 0.35 W:C, in a rockmass of modulus E = 10 GPA,
less than 1 metre of grout bond is required to break a single plain strand cablebolt.
The bond length at which the cablebolt will break is termed the critical bond or
critical embedment length (Section 2.6.1). The charts relating bond strength to
rockmass modulus, grout quality and stress change that are contained in Section
2.6.2 can be used to estimate the critical bond length for plain strand cablebolts.
If a full length cablebolt pulls out of the grout during a pull test, then the grout
is of extremely poor quality, the grout column is full of voids, and/or the rockmass
surrounding the cablebolt has undergone a substantial stress decrease since the
installation of the cablebolts. If the cablebolt does pull out, then it is essential to
determine the cause of the failure, and to install more cablebolts in the area.
In most operational settings, cablebolts with grouted lengths more than 2m
should only be pull tested when serious quality control problems are suspected
such as incomplete hole filling due to grout loss or when serious stress decreases
are suspected. It is not meaningful under normal conditions to perform quality
control field pull tests on design length cablebolts which are fully grouted. The
anchor length being tested (2m) by pulling the cable at the face will almost
certainly be longer than the critical embedment length (required to break the steel
tendon). If these cables pull out at all, then a very serious problem indeed exists.
It is preferable then to periodically install tests cables (not as part of a regular
array but in the same area between operational cablebolts) with all but 250mm of
strand encapsulated in plastic tubing as illustrated in Figure 2.2.5. Grout the hole
normally, leaving a sufficient length of cable extending from the hole for
attachment of a testing jack. The pullout loads obtained can be compared
(normalized with respect to the length of 250mm to obtain bond strength in kN/m)
with the various results shown in Chapter 2.
3.12 Quality Control Improvement
Any problems with the cablebolt installation quality control must be discussed
with the crew(s) and supervisors, and solutions should be found and implemented
as soon as possible.
Descriptions of operational grouting problems and their solutions are given by
Oliver (1992) and Bourchier et al (1992). These papers provide instructive
examples of responses to problems identified in the cablebolt installation process
at specific mine sites.
An interesting case history for quality control improvement at Trout Lake
Mine, Canada was documented by Cluett (1991). The study was initiated when
two large wedge failures occurred in cablebolted cut and fill stopes. When the
quality of the cablebolt installations was investigated, it was found that:
+ The grout was too thin, and the solid cement in the grout mix was settling,
leaving an ungrouted, water filled section at the top of the upholes.
+ The thin grout was escaping from the borehole into joints.
+ The breather tube (6 mm diameter) in long holes (20 m) was being crushed by
the increase in grout pressure in the hole created by trying to pump the grout
back down the breather tube.
+ The old piston pump in use was incapable of grouting the length of boreholes
designed at the specified water:cement ratio.
+ The grout was leaking from the collar of the boreholes.
To solve these problems at the mine, they:
+ Purchased higher pressure rated breather tubes for the piston pump
installations.
+ Purchased a progressing cavity pump
which can pump a thick grout (0.3
W:C) up a long hole (15 m).
+ Examined all of the cablebolt
installations. Where poor quality
cablebolts were suspected, additional
cablebolts were installed.
+ Provided additional training for the
cablebolting crews.
+ Formed a ground control committee
consisting of ground support and
engineering staff, and health and safety
committee members. The members
discuss any problems that arise with
installation and equipment.

+ Cablebolts are now installed soon after the drilling has been completed, so that
the period of time during which the rockmass is unsupported is minimal.
+ There is now a close liaison between the geologists, mining engineers, mine
superintendents, and the mine operators.
+ Grouting guidelines, dependent upon hole length and angle were created. The
piston and progressing cavity grout pumps are both still in use at the mine, but
are only used in situations where it will be possible to completely grout the
borehole. The pumping guidelines are reproduced here in Table 3.12.1 and in
the figure on the preceding page as an example of what can be done to ensure
the correct use of grout pumps at a mine site.
+ The number of boreholes grouted per grout batch is always recorded now.
Table 3.12.1: Grouting guidelines for Trout Lake Mine, Canada (after Cluett, 1991).
Length
of hole
(m)
Angle
from
horiz.
(deg)
Pump Breather
tube
(psi)
Grout
tube
(psi)
Collar
Seal
Grout
mix
(W:C)*
Upholes 14 45 Minepro 450 250 Cement
plug
0.33 to
0.4
9 to 14 45 Minepro None 100 or
250
None 0.3 to
0.33
9 45 Spedel 450 100 or
250
Cement
plug
0.4 to
0.45
9 any Minepro None 75 None 0.3 to
0.33
Spedel 75 75 Cement
plug
0.4
Flat and
down
holes
9 any Minepro None 100 or
250
None 0.3 to
0.33
Spedel None 100 or
250
None 0.4
9 any Minepro None 75 None 0.3 to
0.33
Spedel None 75 None 0.4
* Note that the authors of this handbook recommend a lower bound of 0.35 for
water:cement ratio for most applications (Figure 2.5.9). If grout with W:C = 0.3
to 0.35 can be pumped and placed with confidence and with little difficulty as
in Oliver (1992), then the use of such thick grouts may be warranted.
4 VERIFICATION:
Cablebolt Performance Assessment
4.1 Introduction
The performance of cablebolts in the rockmass should be investigated to ensure
that the support pattern is performing as designed. In some circumstances, the
support pattern may be found to be inadequate for the rockmass and mining
conditions, and in others that the support pattern is over designed.
The extent and mechanism of the rockmass failure, as well as the performance
of the cablebolts, can be assessed visually and through the use of field instruments
in a monitoring program.
+ Observations of failed zones in the rockmass and at the ends of exposed support
elements can be used to determine the rock failure mechanism, to understand
why the cablebolts failed to retain the rockmass, and to provide an estimate of
the load taken by the cablebolts. Detailed field notes and a series of
photographs should be compiled for each working area.
+ Some instruments have been developed in recent years that can augment visual
observations, by viewing data in areas which are not physically accessible.
Borehole cameras allow visual observation of structure or stress induced
spalling along the walls of a borehole. Laser distance meters form the basis of
equipment which, when positioned within a stope, orepass or other non-
accessible mining excavation, will provide a complete 3-dimensional
visualization of the current shape and size of the opening.
+ Other more conventional instruments such as time domain reflectometry
(T.D.R.) devices, extensometers, ground movement monitors, stress cells and
strain gauges can be used to monitor the stress and deformation within the
rockmass. Specialized instruments have been designed to monitor strain along
a cablebolt or the load at the face end of the bolt. When used in combination,
these instruments may record data that can provide a more complete
understanding of the rockmass behaviour and support performance.
+ Computer spreadsheets, graphing tools, and data management software allow
for rapid data reduction and presentation. These tools also facilitate data
analysis and design feedback.

Figure 4.2.1: Visual appearance of failed cablebolts
4.2 Visual Performance Assessment
Visual observations of the state of the rockmass, the final, as-mined boundary
of the excavation and the ends of the cablebolts exposed by rockmass failure are
all important keys to gaining an understanding of the behaviour of the rockmass
and the performance of the cablebolts.
When rockmass failure has occurred, try to assess:
+ The extent of the rockmass failure (% of surface area and depth of failed zone).
+ The appearance of the rockmass. Has there been any spalling or onion
skinning of the rockmass? Are any open joints visible?
+ The approximate size or volume of the blocks that have fallen from the stope
boundaries.
+ The percentage of cablebolts falling into each of the four distinct failure
categories shown in Figure 4.2.1.
+ The length of the cablebolt strands exposed by the failure.
+ Are any plates remaining? If so, where are these plates located, and are they
loose, bent or broken?
Verification: Cablebolt Performance Assessment 363
4.2.1 Remote Visual Data Collection
In recent years, instruments have been developed to allow visual observations
of the rockmass to be made in areas which are not physically accessible. These
instruments include the borehole camera and the remote laser distance meter.
Both of these instruments have been tested extensively in field trials in Canada
by the Noranda Technology Centre. This work has resulted in the production of
readily useable versions of the instruments, which are now commercially
available. A number of mine sites around the world have begun to integrate the
use of the borehole camera and the remote laser distance meter into their regular
monitoring operations.
Borehole Camera
The borehole camera is inserted into a borehole and the head can be rotated to
view and photograph either the wall or along the length of the borehole. Cameras
are available which will fit into a number of different sized boreholes. In a
monitoring program, where regular readings of the instruments will be taken, a
borehole for the camera log should be drilled and left open for the duration of the
work. The collar of the hole should be blocked to prevent mud or cuttings from
falling into the hole.
The borehole camera is being used at the Louvicourt Mine to provide valuable
information about the rockmass behaviour during the initial mine design stage
(Germain, 1995). Before mining of one of the first stope blocks began, a base
survey was made during which the borehole wall was photographed at 0.3 metre
intervals. As mining progressed, borehole camera surveys were conducted on a
regular basis, allowing observation of the development and opening of fractures
and joints within the rockmass.
The borehole camera surveys conducted at Ansil mine (Hutchinson, 1992)
provided some of the most useful information collected during the instrumentation
program, because they provided a visual picture of the location of open fractures
and joints. In addition, horizons within the rockmass where the borehole walls
were spalling due to high stress levels were observed with the camera.
Whenever possible, camera logs should also be made of the boreholes in which
other instruments will be installed. The camera log will provide information about
where pre-existing structures are located along the length of the borehole. This
is particularly important for CSIRO stress cell installation (Maloney, pers.
comm.), and when cablebolt strain gauges are to be used. The location of
movement(s) within the rockmass can be better understood when borehole camera
information is available; from the initial survey of the instrumented hole and from
adjacent holes which are logged regularly with the camera.

Figure 4.2.2: The remote laser distance scanning device: equipment setup and a stope
boundary profile recorded at a mine site (after Miller et al., 1992)
Remote Laser Scanning Device
The remote laser scanning device is mounted on a telescoping arm that can be
extended into the stope, as is shown in Figure 4.2.2, or lowered down a borehole.
Once in the stope, the laser device is rotated in 3 dimensions to accurately survey
the entire stope. The data is recorded directly onto a computer, from where it is
easily loaded into a conventional drawing program such as AutoCad for dataTM
visualization and interpretation. The data that is recorded provides a fairly
accurate picture of the boundary of the stope or excavation that is otherwise
inaccessible for detailed observations of the rockmass.
Figure 4.2.3: Dilution measured with a laser
distance meter at Hemlo Golden Giant Mine
(after Anderson and Grebenc, 1995)
Figure 4.2.4: Final stope boundary measured at Louvicourt Mine (after Germain, 1995)
The adaptation and use of the laser distance device at a number of mine sites
is well documented by Miller et al. (1992). Since the time of that paper, the
device has been tested extensively and improved, and is now commercially
available.
The laser surveying device is used
by the survey crew to measure the
profile of every stope at Hemlo Mine
(Anderson and Grebenc, 1995) and at
Louvicourt mine (Germain, 1995).
The data recorded provides valuable
information about the source and
volume of dilution due to failure of
backfill and waste rock, and indicates
where ore has sloughed into the stope
and where unblasted ore has been left
behind. The results of a laser survey
in a stope which produced a large
volume of waste rock dilution at
Hemlo Mine is shown here. This
information was used to design a
more effective cablebolt pattern for
the adjacent stope which produced
much less dilution (Figure 1.2.2).
When using this equipment, be
aware of any physical and directional
limitations of the hardware: for
example, it is not able to see around
corners. It is important to position the laser device far enough into the stope.

4.3 Monitoring Performance with Instruments
A well designed instrumentation program includes a variety of instruments and
should be directed at answering a specific set of questions. The design of an
instrumentation program aimed at assessing the performance of cablebolts in a
rockmass might include consideration of some of the following points:
+ What is the expected rockmass failure mechanism and which instruments will
provide the information to confirm this hypothesis?
+ Is the rockmass expected to undergo a stress change during the course of the
monitoring program? In instances where the rock slides off the end of the
cablebolts, the bond strength was inadequate. Stress change measurements can
be used to determine if the failure was due to poor installation quality control,
or to a stress decrease, or to a combination of these factors.
+ Is the access for installing the instruments adequate? The read out end of the
instrument must be safely accessible throughout the life of the program, or be
outfitted with a remote read out head that can be protected against damage.
+ Are the methods for protecting the equipment adequate? The instruments,
wiring and any other equipment must be well protected against damage from
falling rock, water, equipment and blast vibrations.
+ Is there sufficient redundancy built into the instrumentation program: both in
the variety of instruments and the number of each kind of instrument? If some
instruments are lost, is it likely that useful information will still be collected by
the remaining instruments?
+ Are enough robust, lower priced instruments used in the program to give good
results? Some instruments are more difficult to install (stress cells) and to
interpret (stress cells and strain gauge instrumented cablebolts) or may be
expensive. These instruments should only be used when the more basic
monitoring of the rockmass behaviour is already assured with the simpler
instruments, when the budget allows for more extensive instrumentation, and
when the information they can provide is worth the cost.
+ Can the purchase or rental of a data logger be justified? A data logger makes
the collection and input of the data very efficient and easy, but adds capital
costs to the monitoring program.
+ How quickly is the information required? Immediate returns may be obtained
from the data collected with instruments that take direct readings, for example
the displacement recorded by an extensometer. This data can be used
immediately for the prediction of stability or impending failure. The full
interpretation of a lot of data from a variety of instruments will take more time.
4.3.1 The Instrument Toolbox
Instrumentation for monitoring the behaviour of the rockmass has been used for
a long time in many mining operations. A good discussion of the wide variety of
instruments, their configuration and their utility is provided by Dunnicliff (1988)
and Franklin and Dusseault (1989).
Case histories illustrating monitoring programs have been compiled by the
C.I.M. (ed. Franklin, 1990) in the Mine Monitoring Manual, and documented by
Goris et al. (1990, 1993), Hutchinson (1992), Kaiser and Maloney (1991),
Maloney et al. (1992), Stillborg (1993), Windsor and Thompson (1993), and
Windsor et al. (1987).
Rockmass Instruments
Different types of instruments record different forms of data. In general
however, instruments record the strain or displacement at a point in the rockmass
or on the support element. General rockmass instruments should provide
information about the:
+ change in the dimensions of the surface of the excavation (convergence
meters),
+ displacement of a point in the rockmass relative to a collar location point
(ground movement monitors, T.D.R., and extensometers), and
+ strain of the walls of a borehole at a specific point (stress change cells).
A single reading from one of these instruments at a specific time is of little use.
Monitoring involves taking a number of readings at different times which will
indicate the relative changes in the parameter measured at that particular point.
When the history of the data is compared with all of the other points, it may be
possible to develop a realistic picture of the behaviour of the rockmass and the
support elements in the area of the study.
As with the cablebolt materials and equipment, there are many manufacturers
and suppliers of rockmass and support instrumentation. There are likely to be
differences in the configuration of each type of instrument and in the ease of
installation and data logging. It is advisable to investigate the range of instruments
available from suppliers, before selecting the equipment for use at the mine site.
The instruments should be robust, accurate, simple and easy to calibrate (Douglas
and Arthur, 1983).
The use of these instruments will not be discussed in detail in this chapter, but
the interested reader is directed to the documents listed above for additional
information.

Figure 4.3.1: Configuration of a spiral strain gauge installed on a cablebolt (after
Choquet and Miller, 1988)
Cablebolt Instruments
The suite of cablebolt instruments currently available on the market is much
more limited than the general rockmass instruments. A number of new
instruments will likely be developed within the next few years, particularly in light
of the increasing use of modified geometry cablebolts in mines, for which there
are no instruments currently available.
Cablebolt instruments currently available provide information about the
average strain along a section of plain strand cablebolt (cablebolt spiral strain
gauge). These data can be used to determine:
+ whether the cablebolts are being loaded close to their breaking strength,
+ if the cablebolts are providing the bond strength used in designed, or
+ if the cablebolt pattern or length should be changed.
The spiral strain gauge can be of any length and will monitor the average strain
in the cablebolt over that length. However, no specific information about the true
distribution of the load applied to the cablebolt is given by the gauges.
The spiral strain gauge consists of a set of plastic sheathed wires that are
wrapped around the cablebolt, following the groove between adjacent,
individual wires, as shown in Figure 4.3.1. The wires are anchored in rubber
plugs that are glued to the cablebolt. As the cablebolt deforms, the wires stretch
within their plastic sheathes, and the resistance of the wire changes. The change
in the wire resistance can then be related to an average load along the cablebolt.
There will be some initial creep of the gauges, for up to 3 months after installation
of the gauge (Choquet, 1993), that should be accounted for in the data reduction.
The application of the correct, small amount of tension to the resistance wires
and the complete waterproofing of the gauges are both very important steps during
the installation of the gauges on the cablebolt. If these two procedures are not well
done, the life of the gauges will be limited. Where available, it is advisable to
purchase the gauges pre-installed at the desired locations on the cablebolt.
Due to the presence of plastic sheathed wires within the grooves between the
individual cablebolt wires, the interaction between the cablebolt and the
surrounding grout will be different than usual. It is thought that the bond strength
and capacity of the cablebolt will be reduced to some extent. In addition should
the grout column move past the instrumented cablebolt for any distance, damage
to the wires is expected to occur. The degree of damage will depend upon the
grout quality, rockmass confinement and any stress change.
A tension gauge can be located at any point along the length of the cablebolt.
The influence of the local geology on the subsequent rockmass movement should
be considered when designing the layout of the gauges. If possible, a borehole
camera should be used to log the position and characteristics of any structures in
the borehole. Knowledge of the pre-existing structural characteristics of the
rockmass can be of great use when interpreting the data recorded by the gauges.
Using this information, the gauges can also be positioned on the cablebolt at the
location of structures that are expected to dilate and load the support.
The spiral strain gauge should be used in conjunction with at least
extensometers and borehole cameras if possible. These other instruments will
provide information about the rockmass failure mechanism and the resultant
source and magnitude of deformations within the rockmass which are loading the
cablebolt. The interpretation of the performance of the cablebolts is very difficult
without rockmass deformation information.
The spiral strain gauge is described in further detail in papers by Windsor et al.
(1987), Choquet and Miller (1988), and Choquet (1993). Other cablebolt and
grout instruments have been developed and investigated by the Commonwealth
Scientific and Industrial Research Organization (CSIRO) in Australia. These
instruments are not commercially available at this time, but the interested reader
is referred to the paper by Windsor (1992).
Several field investigations of cablebolt performance have included strain
gauge instrumented cablebolts (Choquet, 1993; Goris et al, 1990; Hutchinson,
1992; Thibodeau, 1994; and Windsor and Worotnicki, 1986). A review of these
case histories will provide the interested reader with additional information about
the use and interpretation of these gauges.

Figure 4.3.2: Basic diamond-shaped instrument pattern used at Ansil mine
4.3.2 Design of the Instrumentation Program
The selection of instruments for a monitoring program at a particular site will
be governed by budget, past experience, relative cost and applicability of the
instruments, orebody geometry, underground access and the program objectives.
Objective of the Instrumentation Program
The objective of the instrumentation program should be defined before starting
to design the program. The objective of the instrumentation program undertaken
at Ansil mine (Hutchinson, 1992; Hutchinson and Grabinsky, 1992) was to
develop a better understanding of the interaction between the cablebolts and the
rockmass, in order to improve the cablebolt design.
Ideally it should be possible to develop a complete understanding of the
interaction between the rockmass and the cablebolt reinforcement by comparing
the data recorded by different instruments. If this is the aim of the instrumentation
program, then individual groups of the gauges and anchors of the various
instruments should be positioned close to one another in the rockmass. At Ansil
mine the instruments were installed in a diamond pattern in each stope (Figure
4.3.2), which was designed to provide complete coverage of the rockmass at the
centre of the span of each stope, both in terms of spatial coverage and the type of
data collected. The diamond pattern consisted of an extensometer, borehole
camera hole and stress cells in the centre, within the array of cablebolts, some of
which were instrumented with spiral gauges. The monitoring program is discussed
in more detail in Hutchinson and Falmagne (1991) and Hutchinson (1992).
Instrument Cost and Value of the Data
The relative value of the data recorded by each instrument should be assessed,
based on previous experience where possible. The purchase cost of an instrument
is not the only factor in this evaluation. In fact, Dunnicliff (1988) cautions against
the purchase of the cheapest instrument unless it will provide the sensitivity,
accuracy, and longevity required. On the other hand, if it is expensive to purchase
or time consuming to install the instrument or to interpret the data, then the
instrument should be used with caution.
In the early stages of an instrumentation program at a mine site, numerous
inexpensive instruments should be used, if they will provide the information
required. More expensive instruments could then be used when they are required
to answer a specific question arising from the previous instrumentation program.
A ranking of instruments, with respect to the purchase cost, the ease of
installation, and the time required for collecting the data, data reduction and data
interpretation, can be a very useful tool during the design of the instrumentation
program. A ranking based on experience at Ansil mine is shown in Table 4.3.1,
where a data logger collected the extensometer, stress cell and strain gauge data.
A similar ranking should be developed at each site as experience is gained with
different instruments.
The lowest ratings in Table 4.3.1 are for the borehole camera and the
extensometers. If the instrumentation program budget is small, or time for
analysis is limited, then there should be a heavy emphasis on the borehole camera
and extensometer, which are easily installed, read and interpreted. Where
instrumented cablebolts are used, more time has to be spent interpreting and
presenting the data. Stress change cells require a significant amount of time; for
installation and then reducing, presenting and interpreting the data. The cost of
this time should be considered when budgeting the instrumentation program.
The purchase of a data logging device should be evaluated. While the initial
purchase cost may be high, a data logger records data reliably and as frequently
as required. Some of the purchase cost of a data logger is offset by the substantial
reduction in the time spent by personnel reading and inputting data. The
frequency of data collection can be set at any level with a data logger. The data
logger also continues to collect data at times when personnel cannot access the
instrumentation site, such as during production blasts. The data logger selected
for the site must be robust enough to survive the adverse conditions (dust,
moisture and vibrations) found in most underground mines.
Careful field observations, including photographs and detailed notes, can
provide extremely useful information about the deformation and failure of the
rockmass and about the performance of the cablebolts.

Table 4.3.1: Ranking of the instruments used at Ansil mine (after Hutchinson, 1992)
Borehole
camera
Extenso-
meter
Stress
Cell
Spiral
gauge
Unit Cost 1* 2 2 2
Drilling Cost 2 1 3 1
Installation difficulty** 0 1 to 2 3 1
Time for collecting data 2 1*** 1*** 1***
Data reduction difficulty 0 1 3 2
Data presentation difficulty 2 1 3 1
Interpretation difficulty 1 1 3 2
Risk of inaccuracy or misinterpretation 0 1 2 3
Ranking**** 8 9 to 10 20 13
* These costs are for rental of the borehole camera and data logger, and do not consider
the purchase price.
** This ranking assumes that there is easy access to the collar of the camera holes. The
ranking for an extensometer without remote readout head is 1 and with remote readout
head is 2.
*** Data collected with a data logger.
****Increasing ranking indicates greater difficulty or higher cost.
Geometry of the Orebody and Access
The overall geometry of the orebody should be considered in the design of the
instrumentation program:
+ If the boundary of the orebody, and the structure and geology of the rockmass
are fairly regular, then the data recorded by a particular instrument is likely to
be representative of the behaviour of the adjacent rockmass, and as such can be
compared with data collected by adjacent instruments which monitor some
other aspect of the behaviour.
+ Where the geometry, geology or structure of the rockmass is very irregular and
three-dimensional, the data recorded at a point in space is not easily related to
that recorded by a nearby instrument. In this case, more instruments are
required over a smaller area to develop an understanding of the local rockmass
behaviour. A picture of the regional rockmass behaviour will be even more
difficult to obtain.
The design of the instrumentation pattern will also depend upon the
underground access available. Hangingwall, footwall and sill drifts provide
opportunities to install instruments in a number of different patterns.
Redundancy
The instrumentation program must also include some redundant instruments.
All monitoring programs, no matter how well designed, installed or protected will
suffer the loss of some gauges or instruments over the life of the program.
Therefore the number of instruments considered adequate to collect the required
data should be increased by at least 10 to 20%.
During the life of the Ansil monitoring program, 28% of the cablebolt spiral
strain gauges and 14% of the extensometer anchors were lost after providing little
information, due to malfunctioning equipment, operational problems in which
wires were irreparably cut, and due to an unexpected, sudden failure of the
rockmass in one area (Hutchinson and Falmagne, 1991).
Remote Readout
The instruments should be safely and easily accessible throughout the life of the
monitoring program, or else be outfitted with remote readout heads.
Remote readout heads may be required to convert the instrumentation output
into a form that can be recorded by a data logger. This is the case with some
mechanical extensometers for which physical displacement is converted into
electrical signals through the use of potentiometers or voltage transducers. The
heads should be used with caution however, because they can be expensive,
inaccurate, subject to excessive scatter, difficult to install, or may not be designed
to survive the harsh environmental conditions often encountered in mines.
Protection
The instruments, wires and data logger must be protected as much as possible
from damage. Moving equipment can cut or pull down wires and damage the
instruments. Blasting can vibrate instruments excessively, putting them out of
alignment or calibration, or fly-rock can damage the instruments or data logger.
The instruments should be installed from a remote and protected area if
possible, such as from a remote drift or an embayment in the drift wall or from a
cut-out in the rock surface. The instruments and wires should be protected from
damage by very visible signing and physical barricades such as a fence or a plate
over the head of the instrument. Wires should be protected from damage and
vibration: at Ansil mine the wires were placed inside old steel water pipes which
were bolted to the drift back. Slots were cut along the length of the pipes so that
the wires could be inserted easily. The point at which the wires entered the pipes
was protected by short lengths of old air hose slit down one side so that it could
be slipped over the wire (Hutchinson and Falmagne, 1991). Wires could also be
covered with a protective layer of shotcrete.

4.3.3 Installation of the Instruments
The correct installation of the instruments is very important. Problems
encountered during installation or poor procedures are often the source of
subsequent failures of the equipment. Some points to consider when embarking
on the installation process are:
+ Most instruments should be supplied with an installation manual or procedures.
Read the manual first to gain an understanding of the procedures involved.
+ Work through the installation procedures on surface, in a trial run before going
underground. Check that the procedures make sense, and assemble all of the
tools and equipment required. When calibration of the instruments is required,
take the readings at surface if possible.
+ Take a pre-installation reading of all electrical readouts.
+ Measure the length of all holes that are to receive instruments.
+ Keep a detailed record of all observations made during the installation of the
instruments, including any problems encountered. Photos of the rockmass, the
instrument site and the procedure used may be helpful later to evaluate any
problems or subsequent questions. Record a post-installation bench mark
reading for each instrument.
4.3.4 Data Recording
Data should be measured and recorded frequently just after the installation of
the instrument when there are no disturbing influences such as nearby blasting, to
determine if there is any creep or noise of the data readings.
Readings from all of the instruments should be taken just prior to and soon after
any events that are expected to result in changes in the stress or displacement in
the rockmass and support elements.
The data should be plotted in its raw form as soon as possible after the reading
is taken to observe general trends in the data. Take readings as frequently as
required to define trends (Figure 4.3.3). It is never possible to take too many
readings: the data points in a long run of similar results can be discarded, but
missing data points that should have been recorded during an important event
cannot be retrieved after the fact.
Any readings that appear to be anomalous should be taken again as soon as
possible to determine if a mistake has been made or to confirm the value recorded.
Figure 4.3.3: Frequency of readings required to fully define the data set
4.3.5 Data Reduction and Plotting
The data should be reduced and plotted as soon as possible after the readings
are taken. Some suggestions for data reduction and plotting are:
+ Convert the data to the units of interest (mm, MPa, tonnes). The conversion
factors supplied with the equipment should be checked by calibration tests
conducted at the mine site. It is convenient to record the data in a spreadsheet,
where the conversion calculations and plotting parameters are set up.
+ Normalize any calculated data with respect to critical parameters that are
unknown or in which there is little confidence in the true value. For example
the modulus, E, in stress change cell data calculations is not likely to be well
known, but is a direct multiplier of the data. Note this assumption and the
expected value of the factor in the spreadsheet and on the plot.
+ Plot the data on whatever axes are the most convenient and easy to interpret.
The scale of the axes should be established at the start of the monitoring
program, so that the data is always viewed in a similar manner. For example,
average loads recorded by spiral strain gauges on cablebolts should be plotted
on axes scaled to the rupture load of the cablebolt. The data points are usually
plotted with respect to time on the x-axis. Figure 4.3.4 shows some other ways
to plot data that might be easier to interpret in certain situations.
+ Clearly mark any important data on the plots, such as the position of
extensometer points, gauges or cells with respect to the expected stope
boundary, the length of spiral strain gauges on the cablebolts etc. Use a specific
symbol to indicate the time of important events, such as blasts, on the plots.

Figure 4.3.4: Different methods of plotting the data can assist in understanding and
interpreting the results
Figure 4.3.5: Evaluation of rockmass failure
modes from the data distributed along a hole
4.3.6 Data Visualization and Interpretation
The interpretation of the data collected from instruments such as extensometers,
borehole camera logs, stress change cells and cablebolt spiral gauges is an iterative
process, that evolves as more data is collected, as the rockmass failure mode starts
to become clearly understood, and as mining progresses. This process is aided by
effective and timely plotting of the data and visualization of the rockmass and
cablebolt performance.
In general, the changes in extensometer data with time indicate the magnitude
and rate of movement of the rockmass, and the approximate source of the
displacements. Borehole camera logs provide immediate visual evidence of joint
opening, displacement and shear, and borehole spalling in high stress conditions.
The cablebolt spiral strain gauges and the stress change cells indicate the
approximate magnitude of the cablebolt load and the approximate stress change
in the rockmass respectively.
An example of a simple, relatively rapid interpretation of instrumentation data
is given in Figure 4.3.5, which shows several possible rockmass deformations
which could be recorded by extensometers or strain gauged cablebolts.
The interpretation of the results of the instrumentation program for an
evaluation of the interaction between
the cablebolts and the rockmass is
more difficult and time consuming.
The data recorded by the stress
change cells will indicate the
magnitude and orientation of stress
changes. Once joints near the stress
cells open up, the data recorded by
the cells will become fixed in the
direction normal to the joints and the
interpretation of the data will
eventually become invalid as joints
and fractures propagate. The
borehole camera should be used to
log the hole(s), both while the rock is
still intact, and frequently thereafter
as mining causes the rockmass to
deform and the joints to dilate. The
extensometer and instrumented
cablebolts will record meaningful
non-zero data once the individual
rockmass blocks begin to move.

Figure 4.3.6: Data recorded by the instruments
in the hangingwall of the upper horizon stopes
at Ansil mine
Figure 4.3.7: Prediction of cablebolt bond strength (Section 2.6) in a rockmass
undergoing stress change as measured by CSIRO HI cells. The
magnitude of the stress change is represented by the diameter of the
circle, and the orientation of the stress change is plotted as trend/plunge
on the stereonet. (after Hutchinson and Diederichs, 1995)
Data Visualization
The most difficult component of
the analysis of the instrumentation
data is visualizing and understanding
the large volumes of diverse data
collected. The geometry of the
orebody and the location of the
instruments at a particular site may
make the visualisation fairly simple.
However, the compilation of the data
recorded at Ansil mine proved to be
a difficult task, due to the three-
dimensional shape of the
hangingwall, and the large number of
instruments (Hutchinson, 1992). The
best plotting method that was
developed is shown in Figure 4.3.6.
The magnitude of displacement (extensometers) or cablebolt load (spiral gauges)
is indicated by the size of spheres centred on the position of the gauges. Figure
4.3.7 shows the most illustrative way to plot stress change data. These diagrams
were produced for all events of significance at Ansil mine.
Figure 4.3.8: Potential rockmass failure modes to be considered in the interpretation of
data collected from an instrumentation program
Data Interpretation
The interpretation of the results of the instrumentation program should start
with some idea of the rockmass failure mechanism. Three possible modes of
failure are shown in Figure 4.3.8. For each of these failure mechanisms, the
instruments will record different data, thereby either confirming the original
hypothesis about the rockmass failure or leading to some new understanding of the
rockmass behaviour.
Examples of the data that could be recorded by a suite of instruments in a
rockmass failing by wedge failure, by peeling failure or by unravelling failure are
displayed in Figures 4.3.9 and 4.3.10. In these drawings, the tangential stress
change, *F, is oriented parallel to the excavation boundary, while *F , the radialt r
stress change, is perpendicular to the excavation boundary. A black, filled zone
indicates a decrease in compressive stress, while a white zone indicates an increase
in compressive stress. Further details regarding the influence of the formation of
an excavation and possible subsequent rockmass failure on the stress field around
the opening are given by Kaiser et al (1992) and Diederichs et al (1993).
Extensometer data alone will provide some information about the rockmass
failure mechanism (Figure 4.3.11).

Figure 4.3.9: Hypothetical data recorded by rockmass and cablebolt instruments for
different rockmass failure modes: wedge and peeling failure
Figure 4.3.10: Hypothetical data recorded by rockmass and cablebolt instruments for
different rockmass failure modes: unravelling failure
Figure 4.3.11: Interpretation of data recorded by multi-point extensometers (after
Hansmire, 1978)

Figure 4.4.1: Instrumentation Example
4.4 Instrumentation and Failure Analysis
Consider the example stope in
Figure 4.4.1 (top). The inclined, 30m
high hangingwall is in a foliated but
initially unparted schist. Cablebolts are
installed from a auxiliary drive as
shown. A six point extensometer
(EXT), a stress change cell (CSIRO)
and two borehole camera holes (BHC)
are laid out to monitor the hangingwall.
Two BHC logs are required to assess
the persistence of large scale
delamination partings. The logs for the
various instruments for 4 mining stages
(1,2,3,4) are shown (middle).
Information regarding face
displacement (EXT data) and
lamination thickness (BHC data) are
use to calibrate a Model Voussoir beam
(bottom). This analysis is described in
Section 2.18.12. It can be seen that the
initial (back calculated) rockmass
modulus is approximately 50 GPa. The
modulus drops at stage 3 due to blast
damage and an increase in small
partings, although the large scale
delamination does not change. New
lamination partings occur at stage 4. It
can be seen from the Voussoir model
that the hangingwall is reaching critical
displacement and failure is imminent.
The CSIRO cells give information
about stress change, stress, across the
cables. From these readings the change
in cablebolt bond strength is calculated
(Section 2.6.1). By stage 4 the bond
strength has dropped below the critical
value (CBS) and the rock begins to
slide off the cables (see Figure 2.6.3).
Modified geometry cablebolts (Section
2.9) are therefore recommended for
future stopes to prevent beam failure.
Figure 4.5.1: Data example
4.5 Experience: The Best Design Tool
It seems fitting to conclude a discussion of the verification toolbox with a
mention of perhaps the least expensive and most reliable design, verification and
back analysis tool a mine engineer can have - his or her own experience. Since
stopes and mine drifts are typically laid out in often repetitive patterns, experience
gained in every metre of tunnel and with every tonne of ore extracted is directly
applicable to future design (Parker, 1973).
The first step in this process, the maintenance
a tunnel and stope geomechanics database, is
critical. Information on successes as well as
failures can be processed in numerous ways to
optimize design. For example, a simple plot of
intersection span with respect to percentage of
failures in different rockmasses can serve to
prevent the under-design and over-design of
future support systems, thereby increasing safety
and reducing unnecessary costs.
A computer database is a convenient method of maintaining this data (Yazici
and Kaiser, 1992). The information can be kept in a spreadsheet if desired. For
every excavation (drift, intersection, stope) a list of rockmass and environment
information, geometric information and performance data is required for future
analysis. Relationships between rock type, stress level, structural control, local
mine stiffness and failure potential (Kaiser et al., 1992), excavation span and
orientation, mining sequence and support type can be analyzed with respect to
recorded excavation performance (stability, service life, dilution) to determine
critical design parameters in the future. The following is a suggested data list:
Cause
Rockmass: (Section 2.14)
type, strength, structure (joint sets)
quality (Q, RMR, Q, N')
water, moisture
failure potential, local mine stiffness (Kaiser et al., 1992)
Stress: (Section 2.13) overstressed/relaxed
maximum induced stress, pillar stress, depth, stress change (+/-)
Excavation: shape, span, hydraulic radius, type (intersection, drift, stope)
Support: type, age, W:C
Environment: calm/heavy blasting/seismic, extraction ratio
Effect
Excavation: stable/failed/damaged
Failure: Type - slab/wedge/ravelling/cave/burst
failure tonnage, dilution (monitor as in Section 4.2.1)
Support: adequate/damaged/corroded/broken/stripped(cables)

4.6 Application of Performance Assessment
Results to Cablebolt Design
The results of the performance assessment or verification process can lead to
several conclusions:
1) the rockmass is well supported, but the cablebolts are over designed and
therefore more expensive than necessary,
2) the cablebolts are well designed and well installed, or
3) the rockmass is not well supported because the cablebolt design and/or
installation procedure is inadequate, and should be re-evaluated.
Some questions that should be answered in the case of the third result are:
+ Can the installation procedure or equipment be improved, thereby improving
the capacity of the installed cablebolts?
+ Was the capacity and the pattern of cablebolts sufficient to retain the rockmass?
What is the potential rockmass failure mode? Is the best cablebolt type being
used for the failure mode and degree of stress change observed? Are surface
restraint elements required?
+ Can anything in the mining environment be realistically changed to improve the
cablebolt performance? Would the installation of the cablebolts at a different
stage in the mining sequence reduce the effects of destressing at the cablebolt
locations? Are additional drifts required to provide more complete access for
cablebolt installation?
Monitoring programs can be very time consuming and frustrating. The
installation of the instruments, and then the collection, reduction and interpretation
of the data can lead to ambiguous results which only provide a limited
improvement in understanding of the rockmass failure mode or the cablebolt
performance. On the other hand, when monitoring programs do work as designed,
and when the instruments record meaningful data throughout the mining of
adjacent excavations, the results can be extremely useful, leading to improved
design of the cablebolt patterns and a potentially huge savings in money spent on
cablebolt support.
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Index 401
Index
adhesion 79
admixtures 72
accelerators 73
air entrainment 73
corrosion inhibitors 74
expansion agents 74
general guidelines 75
plasticizer 72
thixotropic agents 74
water reducing agents 72
water retention agents 73
applications 36, 284
Cut and Fill 6
drawpoint 6, 276
drifts 5, 264
intersections 6, 264
stope backs 7
stope walls 7
array 34, 36
design 240 - 242
assembly
breather tube 317
cablebolt element 314
grout tube 317
hanger 315, 316
automated cablebolting
machines 341 - 345
cablebolt placement 344
clean up 345
grout mixing 344
grout pumping 344
axial loading 28
axial testing 30, 31
field 31
laboratory 30
barrel and wedge 112-120,125
beam
mandolin bolting 275
support design 274, 382
bent wire hanger 315
birdcaged strand 11, 101
borehole diameter 145
capacity 102, 132
manufacture 125
specifications 131
blast damage, grout 70
bleeding of grout 64
block size 167
influence of scale 175
bond strength 77, 92
adhesion component 79
birdcaged strand 132
bulbed strand 135
buttoned strand 130
bond strength con't
critical 78
dilational component 79
effect, rock modulus 91, 95
effect of rotation 86
effect of rust 85
effect, water:cement 88, 91
epoxy coated 130
epoxy encapsulated 130
frictional component 79
grout shrinkage 87
laboratory testing 92
model 83
nutcaged strand 135
plain strand 77, 127
prediction 92
stress change 93
borehole
cement flow 63
collar sealing 162, 321-324
diameter 88, 141, 145-148
drilling feedback 298
plugs 321 - 324
preparation 298, 311
quality guidelines 303
stiffness 90
Borehole camera 363
boundary crushing 254
breather tube
diameter 148, 158
effect on capacity 346
installation method 15, 142
procedure 317, 327
selection 62, 158
ungrouted 346
brittle rockmass behaviour 174
buckling failure
Euler 263
gravity 265- 274
stress 263
Voussoir 265- 274
bulbed strand 11, 101
capacity 102, 135
manufacture 125
specifications 134
buttoned strand 11
capacity 130
specifications 129
manufacture 125
cablebolt
array length 240 - 242
array spacing 240 - 242
coil diameter 38
composites 259
cablebolt con't
cost 20
cutters 162
density 212, 233, 234, 235
dispensers 159, 160
element 34, 35, 37
element preparation 314
empirical spacing 235, 237
failure observation 362
flutes 38
functions 12, 13
hangers 160, 161
installation options 15, 141
instrumentation 368, 369
lay 37
lay length 28
length 218, 233, 239
length guidelines 212, 215
load transfer 28, 76
manufacture 39
mass 38
observation of failures 362
orientation 36, 124
pattern 5
pigtailed failure 362
placement procedure 308
pull testing 92
purpose 284
pushers 163
ruptured failure 362
sequence 36
shear loading 28, 121 - 123
slings 276
spacers 162
spacing 36, 216, 219
stiffness reduction 104
storage 50
strand construction 39
stress relief 39
stripped failure 362
support pressure 216
tensioning 112 - 120
testing configuration 29, 32
toolbox 11
trucks 163
unravelling failure 362
cablebolt pattern
see Applications
cablebolt placement
downhole 319, 310
uphole, breather tube 318
uphole, grout tube 319
CABLEBOND
bond strength prediction 92
stress change 94, 99

capacity 9, 25
bulbed strand 135
buttoned plain strand 130
epoxy coated 130
fibreglass strand 139
nutcaged strand 135
plain strand 127
quality control effects 346
versus demand 25
capacity considerations
immediate stiffness 26
longevity 26
sensitivity 26
surface retention 26
ultimate ductility 26
ultimate load capacity 26
Cavity monitor 364, 365
cement grout 51 - 54
admixtures 72
blast damage 70
bleeding 64
compressive strength 65
corrosion inhibitors 74
curing time 68
expansion agents 74
field sampling 356
flow 62, 63
hydraulic behaviour 58, 59
load transfer 76
low heat 54
mixing 58, 309, 325
plasticity 58
plasticizer 72
properties 65
pumpability 61
pumping procedure 309
quality 55, 89, 91
sand 54
saturation 58
shear strength 108
shrinkage 87
silica fume 54
stiffness 67, 88
strength 88, 356
sulphate resistance 54
superplasticizers 72
tensile strength 66
thixotropic agents 74
viscosity 58
water:cement 56, 71, 88, 91
water content 56
wet bulk density 57
workability 60
varieties 53
check list, quality 349 - 352
classification, rockmass 177
clean up
grout mixer 326
grout pump 332, 333
cohesion 175
colloidal mixer 151
combination strand 136
compressive strength
grout 65
rock 174
confining stress 175
convergence meters 367
corrosion 45
inhibitors 74
cost
cablebolting 20
instrumentation 371
cotton waste plug 321
crew
payment 279, 281, 283
performance 279
productivity 289
tasks 279
composition 279
training 279, 282, 284
critical bond strength
definition 78
wedge failure 96
curing time, grout 68
Cut and Fill Stopes 5, 6
damage threshold rock 174
data
collection 180, 181
interpretation 377, 379
plotting 375
recording 374
reduction 375
visualization 377, 378
debonding 104
for dynamic loading 104
how to 125
deflection
and stability, Voussoir 268
limiting 268, 269
demand 9, 25
steps 166
versus capacity 25
demand considerations
dilation control 27
displacement 27
gravity loading 27
surface ravelling 27
density, cablebolt 233, 212
design
Cablebolt Cycle 8, 9
cycle 23
grout mix selection 140
limits 217
performance assess 384
design specifications
cablebolt drilling layout 301
cablebolt drilling plan 300
cablebolt installation 307
cablebolt layout 306
material purchasing 296
surveying layout plan 300
deviatoric stress 175
diameter
boreholes 148
breather tubes 148
grout tubes 148
strand 147
dilation
angle, rock 175
control 27
rockmass 257
strength, plain strand 79
dilution 2 - 4, 24
stability graph method 252
DIPS 181
discontinuity strength 180
displacement 27
hard rock 257
limiting 256
rockmass 268, 269
soft rock 257
double plain strand 11
see also Twin strand
downhole preparation 311
drawpoint 5, 276
drifts 5, 264
drilling
equipment 149
feedback 300, 306
procedure 302
drum mixer, selection 151
ductile behaviour 174
ductility, ultimate 26
dynamic loading 259
debonding 104
effect on anchors 120
embedment length 77
empirical design 206 - 220
limits 217
Q 213 - 216
RMR 208 - 212
RQD 207
epoxy coated strand 11, 130
manufacture 125
specifications 128
epoxy encapsulated strand 11
capacity 130
manufacture 125
specifications 128
equipment
borehole collar sealing 162
breather tubes 158
cable dispensers 159, 160
Index 403
equipment con't
cable hangers 160, 161
cablebolt cutters 162
cablebolt pushers 163
cablebolt spacers 162
cablebolt trucks 163
colloidal mixer 151
drilling 149
drum mixer 151
grout mixers 149
grout pumps 149, 155
grout tubes 158
grout tube connectors 162
paddle mixer 152
piston pump 156
portability 150
progress. cavity pump 157
selection 14, 149 - 163
Euler buckling 263
expansive foam plug 324
experience 383
extensometer 367, 381, 382
face pattern, cable array 36
factor of safety 24, 174
failure criterion
Hoek-Brown 175
Mohr-Coulomb 175
failure modes
cablebolts 362
rockmass 379
feedback
cablebolt installation 306
cablebolt layout, plan 300
cablebolt layout, sect. 301
drilling 300, 304, 306
installation observation
report 340
ferruled strand 11, 125
fibreglass strand 138
fish hook hanger 316
fractured ground
grout pumping 331
load transfer 82
frictional strength
interface 85
plain strand 79
rock 175
gravity buckling 265- 274
gravity loading 27
grout -see also Cement grout
flow 142 - 144
mix selection 140
resin 51
shotcrete 51
grout and insert, installation
method 15, 19
grout and retract, installation
method 144
grout collar plug 322
grout mixer
clean up, procedure 326
selection 149, 150
grout mixing 154
colloidal mixer 151
drum mixer 151
paddle mixer 152
grout pump
clean up, procedure 332
installation method 142
selection 142-144,149,155
grout pumping 155 - 157
piston pump 156
procedure 327 - 332
progress. cavity pump 156
grout quality
monitoring 353 - 356
testing 355, 356
visual inspection 353, 354
grout tube 158
connectors 162
diameter 148
install. method 15,17,143
installation procedure 329
procedure 317
selection 62
grouted burlap plug 321
grouting problems 165
twin strand 106
hanger attachment 315, 316
hangingwall
cablebolt capacity 99
stress change 99
hard rock
displacements 257
versus soft rock 12
heavily jointed rock 168
Hemlo Mine 3, 365
high early strength grout 53
Hoek-Brown criterion 174
holding, function 12, 13, 236
hole diameter
bulbed strand 134
buttoned plain strand 129
epoxy coated strand 128
nutcaged strand 133
plain strand 126
hydraulic behaviour, cement
grout 58, 59
implementation
cycle 278
improvement, quality 359
in situ loading 28
influence of scale 175
installation
considerations 51
cycle 278
design specification 307
feedback 289, 308
guidelines 291
instruments 374
observation 348
observation report 340
options 15
problems 291
procedures 305
steps 292
trouble shooting 335
installation accessories
borehole collar sealing 162
cable cutters 162
cable dispensers 159, 160
cablebolt pushers 163
cablebolt spacers 162
cablebolt trucks 163
grout tube connectors 162
hangers 160, 161
installation method 142 - 144
selection 141
instrument, installation 374
instrumentation
borehole camera 363
cablebolt 368, 369
convergence meters 367
costs 371
data acquisition 366
data plotting 375
data recording 374
data reduction 375
extensometer 367
laser distance meter 364
program design 370-373
program objective 370
protection 366, 373
redundancy 366, 373
remote readout 373
rockmass 367
spiral strain gauges 368
stress change cells 367
Tensmegs 368, 369
toolbox 367
interface, frictional strength 85
interpretation, data 377, 379
intersections 5
stability graph method
limitations 247
jack loads 117
jack, tensioning 112,114, 116
jacking procedures 118, 119
joint slip, stress induced 258
jointed rock 168

king wire, definition 37
laboratory testing 92
laser distance meter 364
lay
definition 37
length 38
layouts see Applications
length
cablebolt guidelines 212
empirical 233, 239
rule of thumb 218
limitations, stability graph
method 246, 247
limiting deflection 268, 269
load capacity, ultimate 26
load transfer 76
fractured ground 82
modified strand 103
shear 121 - 123
slab 81
stress change 96
surface anchorage 110
wedge 81
loading
axial 28
combination 28
gravity 27
in situ 28
shear 28
longevity 26
Louvicourt Mine 365
low heat grout 54
Mandolin bolting 275
manufacture
modified strand 125
plain strand 39
massive rock 168
material
handling 295
purchasing 297, 298
quality control 295
material quality control
barrel and wedge 113
cement 55
strand 40, 87
water 55
mechanistic design 253 - 276
mixer - see Grout mixer
selection 150 - 153
mixing, cement grout 58
model
bond strength 83
CABLEBOND 83
modelling, induced stress 172
modified geometry 101
alternatives 125
corrosion 49
load transfer 103
modified Q: Q' 197
Mohr-Coulomb criterion 175
monitoring
grout strength 356
installation practice 348
post install inspection 357
quality control 346
water:cement ratio 355
multiple strand 106
no-support limit 230
no-support span limits 214
nutcaged strand 11, 101
capacity 102, 135
manufacture 125
specifications 133
observation
cablebolt failures 362
grout quality 353
installation practice 348
post installation quality 357
report, installation 340
open breather tube, effect on
capacity 346
open stope design
no-support limit 230
stability graph method 221
orepass support 276
orientation of cables 36, 124
overstress 173
paddle mixer, selection 152
pattern see Applications
payment
cablebolting crew 281
quality control checks 281
peak strength 174
pigtailed cablebolts 362
pipe pumping tests 164
piston pump 156
clean up 333
pumping grout 327, 328
plain strand 11
bond strength 77, 88, 91
capacity 127
dilational strength 79
embedment length 77
frictional strength 79
manufacture 125
specifications 126
plastic analysis 174
plasticity, cement grout 58
plates
load transfer 110
types 111
post installation quality
inspection 357
post peak strength 174
principal stress difference 174
procedure
automated cablebolting
machines 344, 345
borehole collar finish. 321
borehole drilling 302
breather tube attach. 317
cement grout mixing 325
downhole placement 319
downhole preparation 311
element assembly 314
grout mixing 309, 310
grout pumping 327-332
grout pumping in fractured
ground 331, 332
grout tube attachment 317
grouting, breather tube 327
grouting, grout tube 329
hanger attachment 315
hole preparation 311
material handling 297
mixer clean up 326, 332
pipe pumping tests 164
plating cablebolts 118
strand preparation 312
surface fixtures 310, 333
tensioning cablebolts 118
uphole cable placement 318
uphole preparation 311
progressing cavity pump 157
clean up 333
pumping grout 329, 330
protection of instruments 373
pull testing, quality 358
Pump see Grout pump
pump efficiency 62
pumpability, cement grout 61
Q 191 - 196
Q' 197
quality control
borehole preparation 303
breather tube install. 349
cement 55
good practice 290
grout tube installation 351
guidelines 336 - 339
improvement 359, 360
monitoring 290,346, 348
post install inspection 357
pull testing 358
rust of steel strand 85
strand 40, 87
testing 346
water 55
Index 405
quality control guidelines
grout mixing 337
grout pumping 338
surface fixture install. 339
rapid hardening grout 53
ravelling failure 27
rebar 264
reinforcement
cablebolt function 12, 13
function 236
hard rock 257
limiting displacement 256
soft rock 257
relaxation 173
stress 255
remote readout 373
resin collar plug 323
resin grout 51
retention
cablebolt function 12, 13
function 236
surface 26
retracted grout tube,
installation method 15, 18
RMR 186 - 190
rock
block size 167
cohesive strength 175
compressive strength 174
damage threshold 174
dilation angle 175
frictional strength 175
heavily jointed 168
jointed 168
massive 168
modulus 91, 95
shear strength 108, 109
stiffness 90, 91, 179
strength 167, 174, 175, 179
stress 167, 169
yield strength 174
RockMass Rating 186-190
rock mechanics 167
rock modulus 179, 202
buckling 263
beam stability 265
effect on bond
strength 91, 95
rock quality, RQD 182 - 185
Rock Quality Index 191-196
rockburst 174
rockmass
brittle 257
dilation 257
displacement 268, 269
ductile 257
failure modes 379
rockmass con't
hard 257
instrumentation 367
modulus 202 - 205, 265
stiffness 266
strength 200
rockmass classification
basic components 177
comparison of methods 198
N' 222
Q 191 - 196
Q' 197
RMR 186 - 190
rockmass modulus 204
rockmass strength 200
RQD 182 - 185
empirical design 207
rubber collar plug 323
rule of thumb, support 218
ruptured cablebolts 362
rust 85
safety 2, 24
grout mixing 325, 326
guidelines 293
sanded grout 54
saturation, cement grout 58
shear loading 28, 121, 123
shear strength
cement grout 108
grout/rock interface 109
rock 108
shear testing 32, 33
shotcrete 264
shotcrete grout 51
shrinkage of grout 87
silica fume grout 54
slab, load transfer 81
soft rock displacements 257
versus hard rock 12
spacing
cablebolt 235, 237
empirical design 216
rules of thumb 219
span limits for no-support 214
spiral strain gauges 368, 369
spring steel hanger 315
Split set 276
stability 2, 24, 256
unsupported, 208, 230
Voussoir 268
stability graph 221 - 229
calibration 248
case histories 243 - 245
dilution 252
examples 243 - 245
gravity adjustment 228
hydraulic radius 229
stability graph con't
joint orientation 224-227
limitations 231, 246, 247
local conditions 248
mod. stability number 222
N' 222
parametric analysis 249
probabilistic analysis 251
rock stress factor 223
stand-up time 24
standards
strand construction 41
strand performance 43
stereonet 181
stiffness
borehole 90
cement grout 67, 88
debonding 104
displacement 258
immediate 26
reduction 104
rock 90
rockmass 266
rockmass modulus 202
stiffness, E, typical values 179
stope
backs 5, 6
Cut and Fill 5, 6
design 221 - 229
walls 5, 6, 99
strand
breaking load 42
capacity 102
capacity considerations 44
combinations 136
corrosion 45 - 49
debonding 104
diameter 38, 147
elastic modulus 42
elongation 42
fibreglass 138
king wire 37
modified geometry 125
multiple 106
performance 43
preparation 312, 313
proof load 42
proportional limit 42
quality 40, 87
relaxation 42
selection 137
standards 41, 43
stiffness 42
yield strength 42
straps 111
strength
brittle behaviour 174
cement grout 88
cohesion: rock 175
dilation angle 175

strength con't
discontinuity 180
ductile behaviour 174
friction: rock 175
grout testing 356
Hoek-Brown criterion 174
Mohr-Coulomb 175
peak 174
post peak 174
rock 167, 174, 175
UCS: typical values 179
stress
boundary crushing 254
change 93 - 100
confining 175
deviatoric 175
in situ 172
induced 172
induced buckling 263
induced joint slip 258
modelling 172
monitoring 378
on a plane 171
principle 170
relaxation 173, 255
rock 167, 169
shadowing 255
tensor 170
stress change
hangingwall 99
load transfer 96
monitoring 378, 380
re-entrant corners 100
remedial measures 100
wedge 96, 98
stress change cells 367
stripped cablebolts 362
structural data 180
DIPS 181
stereonet 181
sulphate resistant grout 54
superplasticizers 72
support, function 12, 236
support design 23, 25
beam 274, 275, 276
buckling 263
high displacement 258
drift 264
dynamic loading 104, 259
empirical 206, 236
overstressed ground 254
relaxed ground 255
rock wedge 259, 261, 262
rules of thumb 218
seismic 104, 259
stiff support 256
support pressure 216
surface anchorage
barrel and wedge 112-120
surface anchorage con't
load transfer 110
plates 110, 111
straps 111
surface fixture 35
installation 310, 333
surface retention 26
swaged strand 11
swages, manufacture 125
Swellex 259
tendon 35
tensile capacity 77
tensile strength
grout 66
tendon 77
tensioning 35
cablebolts 112 - 120
jack 112, 114
Tensmeg 368, 369
testing
axial 30, 31
configurations 29, 32
constrained 29, 30, 31
field 31
laboratory 30
pipe pumping 164
quality control 346
shear 32, 33
unconstrained 29, 30, 31
training
course contents 283
course frequency 282
Trout Lake Mine 359, 360
tube, selection 62
twin bulbed strand, borehole
diameter 146
twin combination strand 136
twin nutcaged strand, borehole
diameter 146
twin plain strand 11
see also Double strand
capacity 107
grouting problems 106
length 239
spacers 106
spacing 107, 238
UCS, typical rock values 179
see compressive strength
see strength
underhand cut and fill 276
unravelled cablebolts 362
unravelling failure 260, 379
uphole preparation 311
verification
verification con't
example 382
instrumentation 361
observation 361, 362
Victualic pipe collar plug 324
viscosity, cement grout 58
visualization
data from instruments 377
Voussoir buckling 265- 274
Voussoir, support design 274
crit. displacement 268, 382
water, quality 55
water content, definition 56
water:cement ratio
breather tube method 142
bulbed strand 134
buttoned strand 129
definition 56
field testing 355
grout retract method 144
grout tube method 143
nutcaged strand 133
plain strand 126
selection 140
specification 71
visual appearance 353
wedge, rock 379
confined 258
load transfer 81
sliding 260
stability graph method
limitations 247
stress change 96
three-dimensional 262
two-dimensional 261
wet bulk density definition 57
wooden wedge, at collar 322
workability, cement grout 60
yield strength
in situ 174, 254
intact rock 174
Young's modulus
grout 67
rock; typical values 179
rockmass 202
Current Author Contact Information
2007
Dr. D. Jean Hutchinson
Associate Professor
Dept. Geological Science and Geological Engineering
Queen’s University, Kingston, Ontario
Canada K7L 3N6
jhutchin@geol.queensu.ca
Dr. Mark S. Diederichs
Associate Professor
Dept. Geological Science and Geological Engineering
Queen’s University, Kingston, Ontario
Canada K7L 3N6
mdiederi@geol.queensu.ca
www.geol.queensu.ca
www.geoeng.ca
This PDF is provided as a courtesy.
Please reference this material as follows:
Hutchinson, D.J. and Diederichs, M.S. 1996
Cablebolting in Underground Mines.
Vancouver: BiTech. 406p.

Hutchinson and Diederichs
Cablebolting in Underground Mines
1996

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