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DEVELOPMENT OF CORRELATION BETWEEN
ROCK CLASSIFICATION SYSTEM AND MODULUS
OF DEFORMATION
Ph.D. Thesis
KHAWAR MUNIR
2006-PhD-Civil-02
DEPARTMENT OF CIVIL ENGINEERING
UNIVERSITY OF ENGINEERING AND TECHNOLOGY
LAHORE – PAKISTAN
DEVELOPMENT OF CORRELATION BETWEEN
ROCK CLASSIFICATION SYSTEM AND MODULUS
OF DEFORMATION
Thesis Submitted in Partial Fulfilment of the
Requirement for the Degree of Doctorate
in
CIVIL ENGINEERING
By
KHAWAR MUNIR
2006-PhD-CIVIL-02
Approved on: _________________________
Internal Examiner (Research Supervisor):
Name: Prof. Dr. Khalid Farooq
Signature: _______________________
External Examiner:
Name: Dr. Tahir Masood
Signature: ______________________
________________________________
Chairman, Department of
Civil Engineering, UET, Lahore
_______________________________
Dean, Faculty of
Civil Engineering, UET, Lahore
DEPARTMENT OF CIVIL ENGINEERING
UNIVERSITY OF ENGINEERING AND TECHNOLOGY,
LAHORE – PAKISTAN
Dedicated to
My Mother (Late)
i
ACKNOWLEDGEMENTS
All praises and gratitude to the Almighty Allah who has granted me the courage and
tenacity to complete this research thesis.
I would like to express my heartiest gratefulness to Professor Dr. Khalid Farooq, who
supervised this research. I am very much thankful to his encouraging approach, patience,
sound guidance and valuable advice to complete this research. I feel very proud to have
worked under his guidance. Due to his benign support and able coaching I am able to
conclude the research successfully and it is a great privilege to acknowledge his
supervision.
Profound thanks are due to Prof. Dr. A. S. Shakir, Prof. Dr. M. Ilyas and Prof. Dr. Aziz
Akbar for their extreme encouragement and support.
I am very thankful to Dr. Ammad Hassan Khan, Engr. Imtiaz Rasheed and Engr. Hassan
Mujtaba of Civil Engineering Department for their constructive suggestions and
cooperation. I am also very much obliged to Dr. Zia ur Rehman, Transportation
Engineering Department for his proficient advices and encouragement throughout in
research work.
It will be unbecoming on my part if I do not pay credit to staff of Central Material
Testing Laboratories, WAPDA, Lahore. The field testing would have been impossible
without their untiring hard work and extreme dedication. Especially I would like to thank
Mr. Tariq Yousaf, Research Officer, for his tremendous commitment and devotion
during field testing at Basha and Kohala sites.
I take this opportunity to thank the administrative staff of the university in general and
that of the Civil Engineering Department in particular for extending their full cooperation
and help in completing the administrative requirement for this study. Finally I am very
much thankful to my parents, wife, and kids, who always pray for my success and have
been a constant source of love and encouragement.
ii
ABSTRACT
Rock Classification methods are important for the evaluation of different rock
parameters to be adopted for Civil Engineering works. The classification of rock mass
also helps to optimise detailed investigation requirements of a large area. During
preliminary design stage of a project, the classification of rock mass in accordance with
one or more systems can be used to establish engineering characteristics of the rock
mass. This also helps in estimating the strength and deformability of rock mass. A
number of correlations have been developed by various researchers to correlate the rock
mass rating values derived from different systems. Usually, rock mass classification data
are not always available in a format that can immediately be applied to a specific
engineering problem. Therefore, correlations may prove very useful to quickly derive
different design parameters. Furthermore, the availability of the correlations between
classification systems facilitate quick means of verifying resultant rock mass rating
values, without re-calculation of the values.
In this research, four main and well known rock mass classification systems i.e. Rock
Mass Rating (RMR), Tunnel Quality Index (Q System), Rock Structure Rating (RSR)
and Geological Strength Index (GSI) have been applied to the data obtained from Diamer
Basha Dam and Kohala Hydropower Project sites and the rocks have been categorized
according to the numerical values. New correlations among these classification systems
have been developed which can be used for the rocks of northern area of Pakistan.
Generally for a large civil engineering projects; i.e. a tunnel or a dam, modulus of
deformation is required at many locations to understand the behaviour of the rock.
However, sometimes it is not possible to perform several in-situ tests due to time and
funds constraints. Hence it is essential to establish some relationship between rock mass
classifications and modulus of deformation. Another purpose of such studies is to
authenticate the existing correlations being used worldwide. Due to the abovementioned
constraints, it may be uneconomical to conduct tests in all critical areas of a single
project, especially for a large project having highly random rock characteristics. In such
kind of situations, a few large-scale in-situ tests are conducted and correlations are made
between the modulus of deformation values obtained from these tests and different
iii
classification systems. These kinds of correlations can be used for extrapolating the
modulus of deformation which may be a representative of a rock mass condition for
other areas of the project. However the selection of locations of the tests should be done
very carefully.
Empirical correlations between rock mass classification systems and deformation
modulus are useful if a range of in-situ modulus values is desired to be established. Also
the estimated values can be provided for the design. The correlations also indirectly
shape the bases to identify the weak areas in the foundation rock that may affect the
structural behaviour.
In this research, data obtained from Plate Load tests and Flat Jack tests performed at
Diamer Basha Dam and Kohala Hydropower Project have been analyzed to develop the
correlations of modulus of deformation with four rock mass classification systems i.e.
RMR, Q System, RSR and GSI. The Plate Load tests performed at Basha were on large
size plate and deep deformation measurements were made with borehole extensometer
installed underneath the plate.
Based on the rock mass classifications in the four systems, the rock existing at Basha
dam site mainly comprises Fair to Good quality igneous rock while at Kohala site it is
classified as Poor to Fair quality of sedimentary rock units. The correlations developed
among various rock mass classification systems have good regression coefficients
ranging from 0.835 to 0.901 indicating good correlations. During the research the
correlations have been developed between deformation modulus and four (4) rock mass
classification systems. Two different sites of different quality of rocks have yielded
different range of moduli. The correlations developed during present study have been
compared with existing correlations and it has been found that generally these
correlations are in good comparison with the other correlations.
The research will benefit in the design of future hydropower projects of Pakistan in the
region, as the developed correlations may be used to estimate the modulus of
deformation at early design stages.
iv
DEVELOPMENT OF CORRELATION BETWEEN ROCK
CLASSIFICATION SYSTEM AND MODULUS OF DEFORMATION
TABLE OF CONTENTS
Description Page
Acknowledgements i
Abstract ii
Table of Contents iv
List of Symbols & Abbreviations viii
List of Figures x
List of Tables xiii
CHAPTER 1 INTRODUCTION
1.1 General 1
1.2 This Research 2
1.3 Objectives 3
1.4 Methodology 4
1.5 Thesis Overview 5
CHAPTER 2 LITERATURE REVIEW
2.1 Introduction 6
2.2 Rock 6
2.3 Physical Properties of Rock Material 7
2.4 Mechanical Properties of Rock Material 7
2.4.1 Compressive Strength 7
2.4.2 Young’s Modulus and Poisson’s Ratio 8
2.4.3 Tensile Strength 8
2.4.4 Shear Strength 8
2.4.5 Point Load Strength Index 9
2.4.6 Other Mechanical Properties 9
2.5 Relation between Physical And Mechanical Properties 9
v
2.5.1 Relationship between Hardness, Density, and Strength of Rock 9
2.5.2 Effect of Water Content on Strength 9
2.5.3 Relationship between Seismic Velocity and Elastic Modulus 10
2.5.4 Compressive Strength and Modulus 10
2.5.5 Compressive and Tensile Strengths 10
2.6 Modulus of Deformation 10
2.7 In Situ Measurements of Deformation Modulus 11
2.7.1 Plate Loading Test 12
2.7.2 Plate Jacking Tests 15
2.7.3 Flat Jack Test 17
2.7.4 Radial Jacking Tests (Goodman Jack Test) 19
2.7.5 Dilatometer Test 19
2.7.6 Modified Pressure Chamber Test 20
2.7.7 Recessed Circular Plate Test 20
2.8 Rock Mass Classification 21
2.8.1 The Evolution of Rock Mass Classification Systems 22
2.9 Major Rock Classifications Systems 25
2.9.1 The Rock Load Height Classification (Terzaghi, 1946) 25
2.9.2 The Stand-Up Time Classification System (Lauffer, 1958) 26
2.9.3 The Rock Quality Designation Index (Deere et al, 1967) 26
2.9.4 The Rock Structure Rating (Wickham et al, 1972) 28
2.9.5 Geomechanics or Rock Mass Rating System (Bieniawski, 1973) 30
2.9.6 Rock Quality Index (Barton et al, 1974) 33
2.9.7 The Geological Strength Index (Hoek et al, 1995) 39
2.9.8 The Rock Mass Index (RMi) (Palmström 1995) 41
2.9.9 Rock Mass Number and Rock Condition Rating 42
2.9.10 Slope Mass Rating 42
2.10 Comparison of Classification Systems 42
2.11 Correlations between Rock Classifications Systems 43
2.11.1 Significance of Correlations 43
2.11.2 Correlation between RMR and Q System 44
2.11.3 Correlation between RSR and Q System 45
2.11.4 Correlation between GSI and RMR 46
2.11.5 Correlation between RSR and RMR 46
vi
2.12 Correlations Between Rock Classifications Systems And Modulus
Of Deformation
46
2.13 Summary 48
CHAPTER 3 ROCK PROPERTIES OF THE STUDY AREAS
3.1 Introduction 49
3.2 Rock Properties of Diamer Basha Dam Site 49
3.2.1 General Geology 50
3.2.2 Geotechnical Investigations at Basha Dam Site 52
3.2.3 Lab Testing 55
Index Tests 55
UCS, Young’s Modulus and Poisson Ratio 56
Point Load Strength Index Testing 57
Tensile Strength 57
3.2.4 Properties of Rock Mass using RocLab Software 61
3.3 Rock Properties Of Kohala Hydropower Project Site 64
3.3.1 General Geology 65
Sandstone-1 (SS-1) 65
Sandstone-2 (SS-2) 66
Shale 66
3.3.2 Geotechnical Investigations at Kohala Hydropower Project 66
3.3.3 Laboratory Testing 69
3.3.4 Properties of Rock Mass using RocLab 70
3.4 Summary 72
CHAPTER 4 CORRELATIONS BETWEEN ROCK MASS
CLASSIFICATION SYSTEMS
4.1 Introduction 74
4.2 Classification Systems Applied in the Study 74
4.3 Rock Mass Classification of Diamer Basha Dam Site 74
4.3.1 Parametric Study of the Rocks of Basha 76
4.4 Rock Mass Classification of Kohala Hydropower Project Site 88
4.4.1 Parametric Study of the Rocks of Kohala 90
4.5 Correlations between Four Rock Mass Classification Systems 109
4.6 Comparison with Existing Correlations 117
vii
4.7 Summary 120
CHAPTER 5 CORRELATIONS BETWEEN DEFORMATION MODULUS
AND VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
5.1 Introduction 122
5.2 Plate Load Tests At Diamer Basha Dam Site 122
5.2.1 Equipment 124
5.2.2 Methodology 126
5.2.3 Determination of Modulus of Deformation 134
5.2.4 Variation in Modulus of Deformation 139
5.2.5 Average Modulus of Deformation 142
5.3 Plate Load & Flat Jack Tests at Kohala Hydropower Project Site 144
5.3.1 Geology of Adit 2 144
5.3.2 Plate Load Test 145
5.3.3 Flat Jack Tests 149
5.4 Correlations of Modulus Of Deformation 151
5.5 Validation by Artificial Neural Network 157
5.6 Comparison of Correlations with Existing Correlations 161
5.7 Correlations of Modulus of Elasticity and Modulus of Deformation 163
5.8 Summary 166
CHAPTER 6 CONCLUSIONS AND RECOMMENDATIONS
6.1 Introduction 168
6.2 Conclusions 169
6.2 Recommendations for Future Work 170
REFERENCES
APPENDICES
Appendix-A Laboratory Test Results
Appendix-B Details of Engineering Properties Tests
Appendix-C Geological Mapping of Adit 4 of Diamer Basha Dam
Appendix-D Borehole Logs of Kohala Hydropower Project
Appendix-E Plate Load Test Results at Diamer Basha Dam Project
viii
LIST OF SYMBOLS & ABBREVEATIONS
Symbol Description
ANN
Artificial Neural Network
ASTM American Society for Testing and Materials
Ei Modulus of Elasticity
Em or E Modulus of Deformation
GN Gabbronorite
GSI Geological Strength Index
ISRM International Society of Rock Mechanics
IWHR Institute of Water and Hydraulic Research
Ja Joint Alteration number
JC Joint conditions
Jn Joint Set number
JP Jointing parameter joined by empirical relations
Jr Joint Roughness number
Jv Total number of joints in a unit length
Jw Joint Water Reduction factor
L Distance between measuring points
LVDT Linear Variable Differential Transducer
MATLAB Matrix Laboratory
MPBX Multipoint Borehole Extensometer
N Rock Mass Number
NATM New Austrian Tunneling Method
P Pressure in Flat Jack
P Total load on the rigid plate
Q System Tunnel/Rock Quality Index System
Q Pressure on loaded area
R Stress distribution factor
R1 Inside radius of bearing plate
R2 Outside radius of bearing plate
RCC Roller Compacted Concrete
ix
RCR Rock Condition Rating
RMi Rock Mass Index
RMR Rock Mass Rating
RQD Rock Quality Designation
RSR Rock Structure Rating
RTH Rock Testing Handbook
SRF Stress Reduction Factor
SS Sandstone
UCS or σ Uniaxial Compressive Strength
UMA Ultramafic Association
USBR United States Bureau of Reclamation
Vb Block volume.
Wa Average deflection of the plate
Wz Deflection at depth Z
Z Depth beneath center of loaded area
ΔY Deformation between measuring points
𝜐 Poisson’s ratio of the rock
σʹ Major principal stress
x
LIST OF FIGURES
Figure No. Description Page
Figure 2.1: Modulus of elasticity and deformation of rock 11
Figure 2.2: Plate Load test set-up (ASTM 4394-84) 12
Figure 2.3: Close up view of Plate Load test set up 13
Figure 2.4: A typical load-deformation curve for Plate Load test 14
Figure 2.5: Typical set up of Plate Jacking test (ASTM 4395-84) 15
Figure 2.6: Typical anchor location in Plate Load test 16
Figure 2.7: Surface measurements in Flat Jack Test (ASTM D4729-2) 18
Figure 2.8: Flow chart for rock mass characterization and design 23
Figure 2.9: Calculation of RQD 27
Figure 2.10: The Geological Strength Index chart 40
Figure 3.1: Location plan of Diamer Basha Dam and Kohala HPP sites 50
Figure 3.2: A close view of typical Gabbronorite rock piece 51
Figure 3.3: A close view of a UMA rock sample 52
Figure 3.4: Layout plan of Diamer Basha Dam showing Adit 4 and 5 53
Figure 3.5: Portal and inside view of the Adit 4 of Diamer Basha Dam 54
Figure 3.6: Core Examination for Diamer Basha Dam 54
Figure 3.7: Core examination and selection for laboratory testing for Basha site 55
Figure 3.8: Index property tests for Diamer Basha Dam 56
Figure 3.9: Preparation of samples by cutting the cores 58
Figure 3.10: Point Load Strength Index Test 58
Figure 3.11: Preparations for Modulus of Elasticity test 59
Figure 3.12: Unconfined Compression test without strain gauges 59
Figure 3.13: Indirect tensile strength test 60
Figure 3.14: Engineering properties tests performed on cores of Basha site 60
Figure 3.15: A typical plot from RocLab 63
Figure 3.16: The relation between GSI and global strength for Basha 64
Figure 3.17: Layout plan of Kohala Hydropower Project 65
Figure 3.18: Core examination for Kohala Hydropower Project 67
Figure 3.19: Core examination and selection for laboratory testing for Kohala site 68
xi
Figure 3.20: Tests performed on sore samples of Kohala HPP 69
Figure 3.21: GSI vs Global Strength for Kohala HPP 72
Figure 4.1: Typical geological mapping (Ch: 138 – 150) of Adit 4 of Basha site 75
Figure 4.2: Core box of BH 15 (Depth 95 to 100 m) of Kohala site 88
Figure 4.3: Typical borehole logs of BH 11 and 12 showing lithology at Kohala
site
89
Figure 4.4: Frequency of four rock classification systems for Basha site 110
Figure 4.5: Frequency of four rock classification systems for Kohala HPP site 110
Figure 4.6: Comparison of correlation coefficients between various systems 113
Figure 4.7: Correlations between various systems using separate data 114
Figure 4.8: Correlation between Q System and RMR 114
Figure 4.9: Correlation between Q System and RSR 115
Figure 4.10: Correlation between RSR and RMR 115
Figure 4.11: Correlation between RMR and GSI 116
Figure 4.12: Correlation between Q system and GSI 114
Figure 4.13: Comparison of Correlation between Q System and RMR 118
Figure 4.14: Comparison of Correlation between Q System and RSR 118
Figure 4.15: Comparison of Correlation between RMR and RSR 119
Figure 4.16: Comparison of Correlation between RMR and GSI 119
Figure 5.1: View of portal of adit 4 of Diamer Basha Dam site 123
Figure 5.2: Set up for Plate Load test at Diamer Basha site 127
Figure 5.3: Drilling in the floor of the adit to install MPBX 131
Figure 5.4: Rock surface preparation and installation of extensometer 131
Figure 5.5: Different accessories used in Plate Load test 129
Figure 5.6: Flat jack and hydraulic pump 132
Figure 5.7: Installation of plates, flat Jacks and spacers 133
Figure 5.8: Final set up of equipment before load application 134
Figure 5.9: Uniaxial deformations and modulus vs distance in all the tests at
Basha site
139
Figure 5.10: Variation in the modulus of deformation in each of the test at Basha
site
140
Figure 5.11: Variation in modulus of deformation in all the tests at Basha site
(combined)
141
Figure 5.12 Variation in modulus of deformation w.r.to strain at Basha site 142
Figure 5.13: Contours of Modulus of Deformation at Basha site 144
xii
Figure 5.14: Set up of Plate Load test at Kohala HPP site 146
Figure 5.15: Plate Load test in the adit 2 of Kohala Hydropower Project 147
Figure 5.16: Typical load vs deformation curve for Plate Load test at Kohala site 148
Figure 5.17: Geometric terms in Flat Jack test (ASTM 4729-04) 150
Figure 5.18: Execution of Flat Jack test in the Adit 2 of Kohala HPP 151
Figure 5.19: Location of the rock mass classification points 152
Figure 5.20: Correlation between modulus of deformation and RMR 155
Figure 5.21: Correlation between modulus of deformation and Q system 155
Figure 5.22: Correlation between modulus of deformation and RSR 155
Figure 5.23: Correlation between modulus of deformation and GSI 156
Figure 5.24: Output file generated from ANN analysis to validate the modulus
values
158
Figure 5.25: Comparison of modulus of deformation from RMR 159
Figure 5.26: Comparison of modulus of deformation from Q system 159
Figure 5.27: Comparison of modulus of deformation from RSR 160
Figure 5.28: Comparison of modulus of deformation from GSI 160
Figure 5.29: Comparison with Bieniawski’s equation (from RMR) 161
Figure 5.30: Comparison with Barton’s equation (from Q system) 162
Figure 5.31: Comparison with Sarma’s equation (from RSR) 162
Figure 5.32: Comparison with Gokcoeoglu’s equation (from GSI) 163
Figure 5.33: Correlation between Em and Ei 165
Figure 5.34: Correlation between moduli ratio and RMR 165
xiii
LIST OF TABLES
Table No. Description Page
Table 2.1: Major rock mass classification systems 24
Table 2.2: The relationship between RQD and rock mass quality 28
Table 2.3: Rock Structure Rating - Parameter A 29
Table 2.4: Rock Structure Rating - Parameter B 29
Table 2.5: Rock Structure Rating - Parameter C 30
Table 2.6: Input parameters of RMR 32
Table 2.7: Rating adjustment for discontinuity orientations 32
Table 2.8: Rock mass classes determined from total ratings 32
Table 2.9: Meaning of rock mass classes 33
Table 2.10: Rock Quality Designation 35
Table 2.11: Joint set number (Jn) 36
Table 2.12: Joint roughness number (Jr) 36
Table 2.13: Joint alteration number (Ja) 37
Table 2.14: Joint water reduction factor (Jw) 37
Table 2.15: Stress reduction factor(SRF) 37
Table 2.16: Summary of Q system classification 38
Table 2.17: Parameters included in different classification systems 43
Table 2.18: Correlations between RMR and Q system 45
Table 2.19: Correlations between Modulus of Deformation and different rock
mass classification system
47
Table 3.1: Summary of the index properties 56
Table 3.2: Engineering properties of intact rock material for Diamer Basha Dam 61
Table 3.3: Summary of results of rock mass strength for Basha Dam using
RocLab
62
Table 3.4: Rock mass characteristics of selected boreholes of Kohala HPP 68
Table 3.5: Summary of the index properties of Kohala site – mean values 70
Table 3.6: Engineering properties of intact rock material for Kohala site 70
Table 3.7: Summary of results of rock mass strength for Kohala site using
RocLab
71
Table 4.1: Calculation of RMR values for Diamer Basha Dam site 78
xiv
Table 4.2: Calculation of Q index values for Diamer Basha Dam site 81
Table 4.3: Calculation of RSR values for Diamer Basha Dam site 84
Table 4.4: GSI values for Diamer Basha Dam site 86
Table 4.5: Calculation of RMR values for Kohala Hydropower Project site 92
Table 4.5: Calculation of Q index values for Kohala Hydropower Project site 98
Table 4.6: Calculation of RSR values for Kohala Hydropower Project site 103
Table 4.8: GSI values for Kohala Hydropower Project site 106
Table 4.9: Summary of rock mass classifications of Basha and Kohala sites 109
Table 5.1: Locations of the Plate Load tests in Adit 4 of Basha site 124
Table 5.2: Detail of equipment and accessories used in the Plate Load tests at
Basha site
125
Table 5.3: Sequence of applied pressure in Plate Load tests at Basha site 130
Table 5.4: Calculation for the Modulus of Deformation at Basha site 136
Table 5.5: Summary of Modulus of Deformation at Basha site 143
Table 5.6: Details of Plate Load test at Kohala site 146
Table 5.7: Plate Load test results at Kohala site 148
Table 5.8: Calculation for the Modulus of Deformation from Flat Jack tests at
Kohala site
150
Table 5.9: Classifications of rock where the Deformation Modulus was
measured
152
Table 5.10: Modulus of Elasticity, Modulus of Deformation and Moduli Ratio for
Basha site
164
CHAPTER-1
1
INTRODUCTION
1.1 GENERAL
Rock mass classification systems are an integral part of civil engineering, specifically in
the design and construction of underground excavations. Rock mass classification
systems have been developing for over 100 years since Ritter (1879) attempted to build
an empirical approach for tunnel design, especially for determining support requirements
(Hoek, 1994). Laubscher developed the first rock mass classification system for caving
operations in 1975, which was modified by Laubscher et.al., in 1976 (Edelbro, 2003).
Classification methods of rocks are important for the evaluation of different rock
parameters in many Civil Engineering works. The classification also helps to optimize the
need for a detailed investigation of a large area where sometimes the site conditions are
too difficult. In the early design stage of a project the classification of rock mass in one or
more systems can be used to have a preliminary picture of the rock mass and its
characteristics. This also helps in estimating the strength and deformation properties of
rock mass.
A number of correlations have been developed by various researchers to relate the rock
mass rating values derived from different systems to one another. Usually, rock mass
classification data are not always available in a form that may immediately be applied to a
specific engineering problem. Therefore, correlations may be very useful to rapidly derive
different design parameters. Furthermore, the availability of correlation equations
between classification systems facilitates quick means of verifying resultant rock mass
rating values, without re-calculation of the values.
Modulus of deformation is an important parameter required for the design of many civil
engineering structures. It is the property of rock mass usually determined from in situ
tests. Ideally these tests should be conducted at many locations to understand the
behaviour of the rock. However it is not realistic to perform so many in-situ tests due to
time and funds constraints. Hence it is essential to establish a relationship between
modulus of deformation and some important parameters like rock mass classifications.
Another purpose of such studies is that the existing correlations being used in the world
CHAPTER-1 INTRODUCTION
2
can also be verified and improved. In many cases, correlations are developed between the
modulus of deformation values derived large scale in-situ tests and rock mass
classification systems because it may not be economically feasible to conduct tests in all
critical zones for a single project, particularly for large projects founded on highly
variable rock. From such correlations, extrapolation of modulus of deformation values
representative of a wide range of rock mass conditions can be obtained if test locations
are carefully selected.
1.2 THIS RESEARCH
In this research, the rocks of Diamer Basha Dam and Kohala Hydropower Project sites
have been classified into four main and well known rock mass classification systems i.e.
Rock Mass Rating (RMR), Q System, Rock Structure Rating (RSR) and Geological
Strength Index (GSI).
The rocks at both the project sites are different in nature and strength. At Basha, rocks are
strong intrusive igneous rocks while at Kohala, weak sedimentary rocks have been found.
Extensive laboratory testing has been carried out on the rock cores taken from both the
project sites. The parameters obtained from laboratory tests have been used in the rock
mass classifications. The results from the laboratory tests were extrapolated to the rock
mass with the help of RocLab software and different parameters by using the generalized
Hoek Brown Criterion have been computed.
The combination of data of both sites gives a wide range which has a considerable
advantage for regression analyses. Furthermore, the wide rock mass classes also provide
an advantage in the use of the correlations. A total of 143 (48 of Basha and 95 of Kohala)
rating value sets have been used in four classification systems. Using these numerical
values for both the sites, correlations have been developed between classification systems
by regression analyses. These correlations have also been compared with the most
renowned existing correlations being used currently in the world.
The applicability of the classification systems is also discussed in the thesis and some
comments have been included regarding the problems associated with some of the
systems like Q system which is relatively difficult to use having large variations in the
CHAPTER-1 INTRODUCTION
3
input parameters. Therefore, some deviation can be expected in comparison of
correlations involving Q system.
As a part of the research, Plate Load tests and Flat Jack tests performed at Diamer Basha
Dam and Kohala Hydropower Project have been supervised. The Plate Load tests
performed at Basha site were carried out first time in Pakistan where the internal
deformations inside the rock mass were measured with the help of borehole
extensometers. Also, the tests were on bigger size plate i.e. 0.9 m diameter to bear the
heavy pressures during the tests.
Data obtained from the tests, have been analyzed in detail and used to establish the
behaviour of rock mass under applied load cycles and accordingly the variation in
modulus of deformations measured at different points has been examined.
Ninety (90) data sets of deformation modulus and rock mass ratings have been prepared
and plotted to develop the correlations of modulus of deformation with four rock mass
classification systems (RMR, Q System, RSR and GSI). Two different sites of different
quality of rocks have yielded a wide range of moduli which is very good to develop the
correlations. Artificial Neural Network (ANN) has been applied for the validation of
correlations by using the MATLAB software. Also the most famous and reliable
equations have been selected from the literature for comparison with the correlations
developed in this study.
The cores extracted from the holes which were drilled to install extensometers for the
Plate Load tests at Diamer Basha site, were brought to laboratory to conduct the modulus
of elasticity tests. The elastic modulus has also been correlated with in situ modulus of
deformation with a good coefficient of correlation.
The correlations presented in this study have been developed for the first time in Pakistan
and can be used in place of correlations available in the literature.
CHAPTER-1 INTRODUCTION
4
1.3 OBJECTIVES
The principal objectives of the research were set as follows;
i. To develop correlations between different rock classification systems.
ii. To develop correlations between modulus of deformation and rock classification
system.
1.4 METHODOLOGY
The research was carried out through following steps;
• Relevant literature study was conducted throughout the research period using
technical literature existing in libraries and on the internet.
• The author also had a chance to visit the offices of United States Bureau of
Reclamation (USBR) and Institute of Water and Hydraulic Research (IWHR)
China where he discussed the methodology and ideas with eminent experts
regarding the research.
• Extensive laboratory testing on rock core samples of Diamer Basha Dam and
Kohala Hydropower Project sites was carried out.
• Classification of rock masses of both the sites in following four (4) major and
mostly used systems;
– Rock Mass rating (RMR),
– Q System,
– Rock Structure Rating (RSR) and
– Geological Strength Index (GSI)
• Determination of modulus of deformation from in situ tests at both the sites.
• Establishing the correlations between various parameters.
• Validation and comparison of the correlations with the existing renowned
correlations.
CHAPTER-1 INTRODUCTION
5
1.5 THESIS OVERVIEW
Chapter 1 of the thesis presents an introduction to the research topic, a statement as to
why the research was carried out and the study objectives, methodology and scope of
work. Chapter 2 presents a critical review of the literature on the rock mass classification
systems, the evolution of systems. The chapter also includes the philosophy of
quantitative classification systems, the implementation of classification systems in the
civil engineering projects and some merits and demerits of each system. The chapter
presents various in-situ tests to determine the modulus of deformation. Existing
correlations between various rock mass classification systems and their relation with
modulus is also discussed.
In Chapter 3 geological and geotechnical data bases of Diamer Basha and Kohala
Hydropower Projects are presented. Laboratory test results and the formulation of rock
mass parameters are also discussed.
In Chapter 4, details of classification of rock mass of both the sites are presented and new
correlations have been developed by statistical analyses which are also compared with the
most commonly used existing correlations.
Chapter 5 presents the importance of modulus of deformation of rocks and methodology
adopted at both the sites to determine it. The modus operandi and data analysis of the in-
situ tests performed at Basha and Kohala sites to determine the modulus have also been
discussed in the chapter. Correlations have been developed between four (4) rock mass
classification systems and modulus of deformation. Modulus of elasticity which is
determined in the laboratory has also been correlated with in situ modulus of deformation.
Chapter 6 presents the conclusions which have been derived from the research. The need
of additional research on this topic is also presented as recommendations. The research
references and appendices are presented at the end of the thesis.
CHAPTER-2
6
LITERATURE REVIEW
2.1. INTRODUCTION
The development in tunnel engineering and underground structures has raised the
importance of rock mechanics and rock testing. Various new classification systems of
rock mass have been established in the recent past for rock characterization. Some
attempts have also been made to correlate these systems.
The deformability of the rock is the most important and governing parameters among all
and in fact deformability rather than stress is being used for the stability assessment of the
rocks. Some in situ field tests are being used around the world to determine the modulus
of deformation. As an alternative to direct testing methods, the modulus can be estimated
from empirical relationships from the quantitative output of engineering rock mass
classification systems. Several such relationships have been proposed for rock mass
classification systems by many researchers.
This chapter presents the various properties of rocks including physical and mechanical
properties, different classification systems of rock mass and methods to determine the
deformation modulus in the field. Existing correlations among different classification
systems of rock mass and that of classification systems with deformation modulus have
been presented in the chapter too.
2.2. ROCK
In common term, rock usually is a solid mass of natural earth material. All rocks are
composed of minerals. Some rocks are composed of single minerals, but mostly by a
group of minerals. Rock material strength is basically the structural strength of the
mineral composition in a rock material. It is governed by the strength of the minerals
itself as well as the structural bonding of the minerals. Silicate minerals form the largest
group of minerals, and most rocks contain more than 5% of silicates. Some important
rock-forming silicates include the feldspars, quartz, olivines, pyroxenes, amphiboles,
garnets and micas (Zhao, 2010).
CHAPTER-2 LITERATURE REVIEW
7
2.3. PHYSICAL PROPERTIES OF ROCK MATERIAL
Physical properties of rocks are of interest and utility in many fields including
geotechnical engineering. Density, specific gravity, porosity, water absorption and water
content are some of the important physical properties which are determined for an
engineering project.
The mass per unit of volume is termed as density of a material. Density of rock has wide
variations and has significant effect on porosity of rock. The range of most of the rock
densities is between 2,500 and 2,800 kg/m3. Specific gravity is the ratio of rock density to
the density of water and typically ranges from 2.0 to 3.0 for most of the rocks. Porosity
indicates the packing form of the material inside, from densely to lose. Porosity is the
ratio of the volume of non-solid material to the total volume of rock. Therefore, porosity
can be represented in any fraction between 0 and 1. Typically, for solid Granite, the
porosity value is less than 0.01 and for porous sandstone the porosity is 0.5. Water
content can be defined as the measure of the amount of water in a rock material.
Some other crucial physical properties of rocks are hardness, abrasivity and permeability.
Usually all these properties are determined in the laboratory.
2.4. MECHANICAL PROPERTIES OF ROCK MATERIAL
Some of the major mechanical characteristics of rocks are as under;
2.4.1. Compressive Strength
It is one of the most significant property of rocks which is used in design and modeling.
Compressive strength is defined as the material capacity to bear the compressive forces in
axial direction. Commonly, the compressive strength is measured by uniaxial
compressive test or unconfined compressive strength test. Typically compressive strength
of rock is determined from ultimate stress on a stress strain graph.
2.4.2. Young's Modulus of Elasticity and Poisson’s Ratio
Young's Modulus is the modulus of elasticity and defined as the ratio of the rate of
change of stress with strain. However this ratio is for small strains. In other words, it is
CHAPTER-2 LITERATURE REVIEW
8
the measure of the stiffness of a rock material. The modulus can be determined in the
laboratory from the slope of a stress-strain curve which is either obtained in a
compression or in tensile test conducted on rock cores.
Like the strength of the rocks, Young’s modulus also varies widely with rock type. For
very strong and hard rocks, Young’s modulus can be as much as 100 GPa. A few
engineers have established some correlation between Young’s modulus and compressive
strength.
Poisson’s ratio can be defined as the ratio of strain in lateral direction to the strain in axial
direction within a linearly-elastic region. The range of Poisson’s ratio for most of the
rocks lies between 0.15 and 0.4.
2.4.3. Tensile Strength
Tensile strength of rocks may be defined as the maximum tensile stress which the rock
material can resist. This is in fact the ultimate strength in tension of a rock. Rock material
generally has tensile strength in lower range which is due to the presence of micro cracks
in the rock. The micro cracks in a rock may also cause the rock failing abruptly in tension
with some little magnitude of strain.
Tensile strength of rocks can be acquired in the laboratory from several tensile tests.
There are some direct methods and some indirect methods. Brazilian test and flexure test
are most common indirect tests. Due to the complexity in sample preparation, direct test
is not commonly recommended and performed.
2.4.4. Shear Strength
Shear strength of rock material is defined as the strength of rock which resists the
displacements caused by the shear stress. Like soil, the resistance against deformations is
caused by two internal mechanisms i.e. cohesion and internal friction. The cohesion and
friction angle in rocks vary from rock to rock. Direct shear test or triaxial compression
tests are used to determine the shear strength of rock. Generally, the triaxial compression
test is commonly used to determine shear strength parameters. By plotting Mohr circles,
the shear envelope is defined which furnishes the cohesion and internal friction angle
(Zhao, 2010).
CHAPTER-2 LITERATURE REVIEW
9
2.4.5. Point Load Strength Index
Point load test is an index test for the strength of rock material. It is a simple test which
gives the standard index of point load. The index is calculated from the point load at
breakage of sample and the size of the specimen. Size correction is also applied for a 50
mm equivalent core diameter.
2.4.6. Other Mechanical Properties
Some other mechanical properties which are required sometimes in some special
problems are as under;
Fracture toughness
Brittleness
Indentation
Swelling
2.5. RELATION BETWEEN PHYSICAL AND MECHANICAL PROPERTIES
2.5.1. Relationship between Hardness, Density, and Strength of Rock
Rebound hardness by Schmidt hammer is usually measured at initial stages of field
investigations. Sometimes to estimate uniaxial compressive strength of the rock, the
hardness index is used. The relationship between compressive strength and hardness of
rock is influenced by the density of the material.
2.5.2. Effect of Water Content on Strength
Research shows that water content has significant effect on rock strength. When rock is in
saturated or wet condition, the uniaxial compressive strength may be decreased in
comparison to the rock strength in dry state.
2.5.3. Relationship between Seismic Velocity and Elastic Modulus
Seismic wave velocity which is determined by geophysical tests indicates physical extent
of the elastic properties of rock. It is used to have a reasonable estimation of the elastic
CHAPTER-2 LITERATURE REVIEW
10
modulus of the rock. From the theory of elasticity, P-wave velocity is related to the
density of the material and resultantly elastic modulus of rock can be determined.
2.5.4. Compressive Strength and Modulus
Generally a stronger rock material is also firm and stiff. Therefore, higher values of
elastic modulus mean higher strength of rock material. Many reasonable correlations have
been developed between compressive strength of a rock and its modulus.
2.5.5. Compressive and Tensile Strengths
Typically, tensile strength is about 10% to 12% of the uniaxial compressive strength of a
rock material. Therefore rock failure may occur easily under tension. That is why, while
designing, the rock should not be subjected to high tensile stresses.
2.6. MODULUS OF DEFORMATION
The full description of deformability of the rock should include not only the elastic
parameters i.e. modulus of elasticity and Poisson’s ratio but also the permanent
deformation with any applied level of stress. The stress to permanent deformation ratio
observed on releasing that stress to zero is called the modulus of deformation (Goodman,
1989). The static modulus of deformation is one of the parameters which best represent
the mechanical behavior of a rock and of a rock mass.
Modulus of elasticity is the property of intact rock usually measured in the laboratory
while deformation modulus is the property of rock mass which is measured by field tests.
Both these moduli are statically determined. Figure 2.1 shows the elastic modulus and
yield function which is incorporated in deformation modulus.
The deformation modulus is one of the parameters which correspond to the mechanical
behavior of a rock mass. The parameter is very important in underground excavations and
tunneling. For this reason, mostly the boundary element and finite element analyses are
based on deformation modulus to study the behaviour of the stress and strain distribution
around any underground excavations. The modulus of deformation is therefore a keystone
of many geomechanical analyses (Palmström, 2001).
CHAPTER-2 LITERATURE REVIEW
11
Figure 2.1: Modulus of elasticity and deformation of rock (Goodman, 1989)
Several investigations have lead to the fact that the modulus of deformation determined in
the field is not constant and depend on the condition of stress. The modulus is generally
higher in rocks subjected to high stresses than in rock masses which have low stresses.
Furthermore, higher stress occurs in better rock mass quality. Different methods and
equipment used to determine the value of design modulus of rock mass may give different
conclusion (Palmström, 2001).
2.7. IN SITU MEASUREMENTS OF DEFORMATION MODULUS
All the field tests to determine deformation modulus are difficult to carry out and also
these are very expensive. The tests are mostly conducted in test Adits or chambers
specially excavated for the tests by conventional drill and blast method. Usually the width
of such Adits is around 2 m and a height less than 3 m. However, the dimension of such
Adits depends upon local conditions as well. Initial preparations for the test at the site are
very lengthy. Another difficult feature in such tests is the interpretation of in situ data,
which requires knowledge and expertise.
CHAPTER-2 LITERATURE REVIEW
12
Presently, the following types of in situ tests are commonly used to determine the
deformation modulus:
2.7.1. Plate Loading Test
The method is based upon the measurement of the deformations at the rock surface which
is subjected to loading as shown in Figure 2.2.
Figure 2.2: Plate Load test set-up (ASTM Designation D4394-84)
This can be easily arranged in an Adit or underground chamber horizontally or vertically.
The site for the tests is selected carefully and the locations having representative rock
mass quality should be selected. Fractured zones or rock mass features having faults,
folds or cavities etc should not be chosen. The rock surface is leveled and a thick mortar
pad is applied on the surface. The plate size may vary from 30 cm to 100 cm. Plate load
test may be a rigid or a flexible. The plate is called a rigid plate when deflection from
CHAPTER-2 LITERATURE REVIEW
13
center to edge of plate is less than 0.0001 inch or 0.0025 mm, when maximum load is
applied.
The depth of the rock volume which is affected by the loading is proportional to the plate
diameter. As the application of very high pressure may not be possible, it is advisable to
use a plate dia upon which the desired pressure can be applied easily. The load can be
applied against the walls of the Adit by means of hydraulic cylinders or screw jacks. Flat
Jacks may also be also used to transfer the load smoothly to the rock surface. The
displacements must be measured at several places on the bearing plate to consider the
circular affect of plate and any possible bending. The deformations are usually measured
by dial gauges which are carefully mounted on the plate as shown in Figure 2.3.
Figure 2.3: Close up view of one end of Plate Load test set up (Goodman, 1989)
Generally, five pressure cycles up to peak pressure, each in ten increments at the rate of 1
minute per increment, are sufficient. If possible, the middle cycle should be
approximately at the level of design load and the upper cycle approximately two times of
the design load. All the cycles need not to be uniformly spaced. However, the unloading
phase of each cycle should be at the same rate as the loading rate. Take deflection
readings from dial gauges after each load increment and decrement. Maintain the peak
and zero pressures for at least 10 min for each cycle, and deflection readings taken at 5-
minutes intervals.
CHAPTER-2 LITERATURE REVIEW
14
If required, both instantaneous deformation and the creep can be determined from this test
method. The modulus, E or Em is calculated by the formula as per ASTM Designation
D4394-88, as follows:
E = 1 − ν2 P
2RWa (2.1)
Where
ν = Poisson’s ratio of the rock,
P = Total load on the rigid plate in lbs or kN,
R = Radius of the plate in inches or mm,
Wa = Average deflection of the plate in inches or mm,
The data can be plotted and a report is prepared including all the results. A typical five (5)
cycle curve is presented in Figure 2.4.
Figure 2.4: A typical load-displacement curve for Plate Load test
2.7.2. Plate Jacking Tests
The only difference between Plate Loading and Jacking test is that in Plate Loading test
only surface deformations are measured while in Plate Jacking test, the rock
displacements are measured in boreholes already drilled behind each loaded area with the
CHAPTER-2 LITERATURE REVIEW
15
help of some deep deformation measuring device like borehole extensometer. However in
several literatures, Plate Jacking is also termed as Plate Loading test. A sketch showing
the typical test set up is presented in Figure 2.5.
Figure 2.5: Typical set up of Plate Jacking test (ASTM Designation D4395-84)
The deformations behind the loaded area in both the boreholes on opposite faces are
measured by a reliable multiple position borehole extensometer (MPBX). For
measurements on the surface, dial gages, or linear variable differential transducers
(LVDTs) may be used, if required. The recommended accuracy by ASTM is at least
0.0001 in. (0.0025 mm), and a sensitivity of at least 0.00005 in. (0.0013 mm). More
accuracy is required in hard rocks. The measurements of deformation within the rock
mass should be taken along a line within 5° of the direction of loading. Furthermore, the
line should be near to the centre point of the bearing pad. The holes for extensometer
would depend upon the size of extensometer but these should be as small as feasible. The
holes should drilled by diamond-rotary on exactly opposing surfaces and should be
continuously cored and logged. Select the location of each measurement point in the hole
by examining the rock core. The borehole can be inspected with a borescope or other
CHAPTER-2 LITERATURE REVIEW
16
suitable device if available. At least two measuring points within the rock surface should
be placed equal to Flat Jack diameter. The deepest two measuring points of extensometer
should be placed at least six Flat Jack diameters from the bearing surface so that they
should be outside the theoretical zone of influence. However, the arrangements depend
upon the specific geologic conditions as shown in Figure 2.6. The extensometer wires are
carefully extended out from the hole and from the side of the bearing pad.
Figure 2.6: Typical anchor locations in a Plate Load test (RTH-365-80)
Cores recovered, if any, should be logged and tested for rock quality designation (RQD),
fracture spacing, strength, and deformation.
The modulus of deformation E or Em, can be calculated from the deflection at a point
beneath the center of an annularly loaded area within the rock mass as follows:
𝐸 =2𝑄 1 − 𝜈2
𝑊𝑧 𝑅2
2 + 𝑍2 1 2 − 𝑅12 + 𝑍2 1 2
+𝑍2𝑄 1+𝜈
𝑊𝑧 𝑅1
2 + 𝑍2 −1 2 − 𝑅22 + 𝑍2 −1 2 (2.2)
CHAPTER-2 LITERATURE REVIEW
17
Where;
ν = Poisson’s ratio of the rock,
Q = pressure on loaded area, (MPa),
Z = depth beneath center of loaded area, (mm),
Wz = deflection at depth z, (mm).
R2 = outside radius of bearing plate, (mm), and
R1 = inside radius of bearing plate, (mm).
2.7.3. Flat Jack Test
In a Flat Jack test, apart from modulus of deformation, stresses at the rock surface and the
long-term deformational properties e.g. creep can also be determined. The test measures
the average stress perpendicular to the surface of the test Adit or chamber. In this test, the
in situ stresses in the rock mass are relieved by cutting a slot of standard dimension into
the rock which is perpendicular to the surface of the test Adit. The deformation is
measured which is caused by this stress relief. A hydraulic Flat Jack is inserted into this
slot and the pressure is applied to this Flat Jack until the early measured displacement is
canceled. The stress reapplied is approximately equal to the stress in the rock mass in a
direction perpendicular to the plane of the jack at the test location. To determine the
deformational properties of the rock mass, loading of the Flat Jack is carried out
incrementally and the resultant deformations are measured. The most accurate direction
of stress determination in a Flat Jack test is parallel to the longitudinal axis of the Adit.
The reason is that, in this direction, the stress is the least affected by the presence of the
opening. A sketch showing the recommended array measurement is presented as Figure
2.7.
CHAPTER-2 LITERATURE REVIEW
18
Figure 2.7: Surface measurements in Flat Jack test (ASTM Designation D4729-2)
When deformation is measured on one side of the slot, the modulus, E, is calculated using
the following equation.
E = PLR 2π∆Y (2.3)
Where:
P = pressure in Flat Jack, lbs/in2 (MPa),
L = distance between measuring points, inch (mm),
R = stress distribution factor, and
ΔY = deformation between measuring points, inch (mm).
When deformation measurements are taken across the slot, Eq. (2.3) is rearranged to
solve for the modulus, E:
E = KP/∆Y (2.4)
CHAPTER-2 LITERATURE REVIEW
19
Where:
P = pressure in Flat Jack, lbs/in2 (MPa),
ΔY = deformation between measuring points, in. (mm),
and
K = coefficient dependent on test geometry
2.7.4. Radial Jacking Tests (Goodman Jack Test)
This test is also known as Goodman Jack test. The test is carried out by a tool which is
used in borehole to estimate the in situ deformability of rock mass. It is designed to be
used in 76 mm boreholes by loading a test chamber in radial direction. The borehole of
circular cross section is uniformly loaded. The Goodman Jack is attached to a drill rod
and inserted into the borehole, along with its hydraulic accessories and signal cable.
The jack has two curved rigid bearing plates at 90o. The plates can be forced apart inside
a bore hole of NX size by pistons. For measuring the displacements, two transducers are
mounted at either end of the 20 cm long bearing plates (Goodman, 1989).
To conduct a test, a test chamber is excavated in circular shape and a uniformly
distributed pressure is applied to the chamber walls by means of Flat Jacks located on a
reaction frame. The extensometers placed in boreholes perpendicular to the chamber
walls record the deformations. Pressure is measured by a hydraulic transducer. During the
test, the pressure is applied in cycles with increments and deformations are recorded at
each increment. The deformation modulus is then calculated.
2.7.5. Dilatometer Test
The dilatometer test methodology is based on the theory of elasticity according to which
the rock mass is an elastic, isotropic and homogeneous medium. The dilatometer is one of
the most adaptable instruments used for determining the modulus of deformation in the
field. Through dilatometer the hydraulic pressure is applied on the rock mass through a
flexible membrane in boreholes. During test it is possible to define the deformational
behaviour of rock mass with reference to the relationship between the pressure and
CHAPTER-2 LITERATURE REVIEW
20
deformation. Therefore, behavior of rock mass for under loading and unloading
conditions can be evaluated (Morteza et al., 2010)
Usually the probe of dilatometer comprises a metal cylinder with displacement sensors.
The pressure is applied to the wall of the borehole uniformly which causes the borehole to
expand outwards. The displacement transducers in borehole are in contact with the rock
surface and measure its deformation.
The interpretation of the test results is rather difficult due to the variation of behavior of
rock mass during testing. Generally, the modulus increases as the applied pressure is
increased. The reason is the closure of cracks or joints in the rock mass under pressure,
making the material more stiff at higher stresses (Morteza, 2010).
2.7.6. Modified Pressure Chamber Test
The rock mass deformability is measured by pressurizing the cylindrical walls of a
chamber or Adit hydraulically and the deformations are measured subsequently. The
modulus of deformation is calculated then. As the load is applied to the large volume of
the rock mass, the results may be declared as true representatives of the rock mass by
considering the effect of discontinuities. The anisotropy of modulus can also be measured
by this method. The results are used for designing the dam foundation, tunnel lining and
pressure shaft (RTH-366-89, 1993).
The test chamber is carefully selected and lined with concrete. All the measuring
equipment is carefully located. Reaction frame is assembled and all the accessories are
checked before loading the chamber. Loading and corresponding deformations are then
recorded to evaluate the modulus.
2.7.7. Recessed Circular Plate Test
In this test the flat surface of the floor of a borehole is loaded with a plate and the
resultant deformations are measured. Elastic modulus, deformation modulus and creep
can be measured from this method. Many horizons can be tested in the same set up by
using a large diameter drill. The direction of loading should coincide with the axis of the
hole. Diameter of the hole is usually kept 860 mm.
CHAPTER-2 LITERATURE REVIEW
21
2.8. ROCK MASS CLASSIFICATION
In feasibility and initial design stage of a project, when there is very less detailed
information on the rock mass and its properties, the rock mass classification schemes can
be of used with considerable advantage. Also, we may use the classification systems as a
check-list to make certain that all pertinent information has been incorporated. One of the
major benefits is that one or more rock mass classification systems can be used to have a
scenario of the rock mass to provide estimation of support requirements, strength and
deformability of the rock mass (Hoek, 2007).
Different rock mass classification systems have been developed since last 100 years. In
1879, Ritter attempted to develop an approach which was purely empirical for support
requirements in tunnel designing (Hoek et al., 1995).
The purposes in application of a rock mass classification system are followings;
(Bieniawski, 1976):
a. To place a rock mass into grouping having same behavior,
b. To provide a foundation for accepting the properties of each group,
c. To assist in the stages of excavations in rock by providing quantitative data which is
usually required for the solution of engineering problems,
d. To provide a common base for successful communication between all professionals
connected to a geotechnical project.
Therefore, a classification system should have the following qualities to fulfil these
objectives:
i. Simple and understandable,
ii. All terms clear and terminology used should be widely acceptable,
iii. Should have only the most significant properties of rock mass,
iv. Should have the parameters which can be determined by quick and cheap tests in
the field,
v. Should be general in use so that a particular rock mass will have the same basic
classification for various structures such as tunnels, slopes and foundations etc.
CHAPTER-2 LITERATURE REVIEW
22
In any quantitative classification system, minimum rating is assigned to the poorest rock mass
2.8.1. The Evolution of Rock Mass Classification Systems
Rock mass classification systems have been a vital part of civil engineering, specifically
in the design and construction of underground structures. An analysis of geotechnical
literature points out that several systems for rock mass classification have been developed
since 1940s. However, there were some limitations which remained associated with the
development of a satisfactory rock mass classification system. These limitations were
identified by Bieniawski (1973), and included that the classification systems were
impractical, too general, entirely based on rock characteristics and did not emphasize the
properties of rock mass. The development of the various rock mass rating systems has
been that the systems were started with classification for use in design. Later, by different
modifications, the systems have been applied for characterization during site
investigations. The term classification should preferably be used for the design of the
excavation as the rating systems are design tools also (Roshoff, 2002).
The characterization / classification of rock mass is based on data from geological
investigations, rock mechanics, hydrogeology and geophysics collected from the field as
well as from laboratory tests. The volume of input data will increase from the beginning
of a site investigation to the final construction stage. The data points usually concentrate
along boreholes and on surface mapping locations inside tunnels or Adits.
The behavior of rock changes drastically from continuous elastic of intact rock materials
to discontinuous running of highly fractured rock masses. The existence of rock joints and
other discontinuities plays important role in governing the behavior of the rock mass
properties.
It is important to understand that requirements for rock classification for design and
characterization are different. Due to this reason, different treatment of parameter values
and their weights to the overall rating indices is required. The subject is relatively new
and certain level of risks about the validity and consistency of the methodology and
results must be taken in these regards. A flow chart for rock mass characterization and
classification from Palmström et al., 2001 is presented in Figure 2.8.
CHAPTER-2 LITERATURE REVIEW
23
Figure 2.8: Flow chart for rock mass characterization and design (Palmstörm et al., 2003)
One of the advantages of the empirical approach is that it is handy to represent the
inconsistency of the properties of rock mass. This can be achieved by statistically treating
the ratings and the mechanical properties derived from the characterization for determine
ranges of variation and spatial trends, it is important that enough data from surface and
underground mapping and experimental measurement (both geological, geophysical and
mechanical) are gathered to achieve acceptable consistency of the results so that a too
pessimistic or optimistic evaluation of the rock conditions is avoided.
Apart from all limitations, about twelve classification systems which were developed
between 1946 and 2002, are being used for rock mass classifications successfully. A list
of major rock mass classification systems is presented in Table 2.1.
Possible
Behaviour of
Ground
Giving Values
to the Various
Rock Mass
Features
Other
Calculations
Numerical
Modelling
Classification
Systems
Project Related
Features
Groundwater
Rock Stresses
Characterization Applications Field Observations or
Measurements
Rock
Mass
Density and
Pattern of Joints
Intact Rock
Characteristics
Joint
Characteristics
CHAPTER-2 LITERATURE REVIEW
24
Table 2.1: Major rock mass classification systems (Palmström, 1995, 2003)
Name of
Classification
Author and
First version
Country
of origin.
Applications Form and
Type*)
Remarks
Rock Load Theory Terzhagi, 1946 USA Tunnels with steel
supports
Descriptive F
Behaviour F,
Functional T
Unsuitable for
modern
tunnelling
Stand up time Lauffer, 1958 Austria Tunneling Descriptive F,
General T
Conservative
NATM Rabcewicz,
1964/65 and
1975
Austria Tunneling in
incompetent
(overstressed)
ground
Descriptive F
Behaviour F,
Tunneling
Utilized in
squeezing
ground
conditions
RQD Deere et al.,
1966
USA Core Logging,
tunneling
Numerical F,
General T
Sensitive to
orientation
effects
A recommended
rock classification
for rock
mechanical
purposes
Patching and
Coates, 1968
For input in rock
mechanics
Descriptive F,
General T
The Unified
classification of
soils and rocks
Deere et al.,
1969
USA Based on particles
and blocks for
communication
Descriptive F,
General
i) RSR concept Wickham et. al.,
1972
USA Tunnels with steel
support
Numerical F,
Functional T
Not useful with
steel fiber
shotcrete
RMR System
(CSIR)
Bieniawski 1974 South
Africa
Tunnels, mines,
foundations etc.
Numerical F,
Functional T
Unpublished
base case
records
Q-system Barton et al.
1974
Norway Tunnels, large
chambers
Numerical F,
Functional T
Mining RMR Laubscher, 1975 Mining Numerical F,
Functional T
The typological
classification
Matula and
Holzer, 1978
For use in
communication
Descriptive F,
General T
Not presented in
this report
ii) The Unified
Rock Classification
System (URCSS)
Williamson,
1980
USA For use in
communication
Descriptive F,
General T
Basic geotechnical
description (BGD)
ISRM, 1981 - For general use Descriptive F,
General T
Rock mass strength
(RMS)
Stille et al.,
1982
Sweden Numerical F,
Functional T
Modified RMR
Modified basic
RMR (MBR)
Cummings et
al., 1982
Mining Numerical F,
Functional T
Simplified rock
mass rating
Brook and
Dhamaratne,
1985
Mines and tunnels Numerical F,
Functional T
Modified RMR
and MRMR
Slope mass rating Romana, 1985 Spain Slopes Numerical F,
Functional T
Ramamurthy /
Arora
Ramanurthy and
Arora, 1993
India Fr intact and jointed
rocks
Numerical F,
Functional T
Modified Deere
and Miller
approach
Geological
Strength Index-GSI
Hoek et al.,
1995
Mines, tunnels Numerical F,
Functional T
Rock Mass Goel et al., 1995 India Numerical F, Stress free Q
Continued...
CHAPTER-2 LITERATURE REVIEW
25
Name of
Classification
Author and
First version
Country
of origin.
Applications Form and
Type*)
Remarks
Number N Functional T System
Rock Mass Index
RMi
Arild
Palmstorm,
1995
Norway Rock Engineering,
Communication &
Characterization
Numerical F,
Functional T
*) Definition of the following expressions
Descriptive F = Descriptive Form: the input to the system is mainly based on descriptions
Numerical F = Numerical Form: the input parameters are given numerical ratings according to their characters
Behaviouristic F = Behaviouristic Form: the input is based on the behaviour of the rock mass in a tunnel
General T = General Type: the system is worked out to serve as a general characterization
Functional T = Functional Type: the system is structured for a special application (for example for rock support)
2.9. MAJOR ROCK CLASSIFICATIONS SYSTEMS
2.9.1. The Rock Load Height Classification (Terzaghi, 1946)
In 1946, Terzaghi worked on the use of rock mass classification for design of support
system for tunnel. The system was descriptive in nature and was focused on the properties
which dominate behaviour of the rock mass when gravity is the dominant force.
Terzaghi’s Classification System comprised rock mass descriptors as follows;
– Intact rock
– Stratified Rock
– Moderately Jointed Rock
– Blocky and Seamy Rock
– Crushed but Chemically Intact Rock
– Squeezing Rock
– Swelling Rock
The estimated support pressure has a wide range to work with squeezing and swelling
rock conditions for a significant application. However it gives over-estimates for tunnels
having diameter of more than 6 m (Singh and Goel, 1999). According to Bieniawski
(1973), the Rock Load Height Classification System is only applicable to the tunnels with
steel supports and not applicable to modern tunneling methods which use shotcrete and
rock bolts etc.
CHAPTER-2 LITERATURE REVIEW
26
2.9.2. The Stand-Up Time Classification System (Lauffer, 1958)
The system is a tunneling-based classification system. The system suggested that for an
unsupported span, the stand-up time is linked to the quality of the rock mass where the
portion of the tunnel is excavated. The stand-up time concept is that an increase in the
span of the tunnel leads to a significant reduction in the time available for the installation
of support. Therefore, the concept of standup time is related to the excavation size, i.e. for
larger excavations, time available prior to failure will be greater. This system is most
appropriate in soft ground e.g. shale, mudstone and phyllite or in highly broken rock
where squeezing and swelling of ground can create stability problems. In hard rock
excavations as the stability is not time dependent, therefore this system is not truly valid.
The Stand-Up Time Classification System has been modified and now it is used as a part
in a new tunneling approach called as the New Austrian Tunneling Method (NATM)
(Pacher et al., 1974).
Bieniawski (1973) considered the Stand-Up Time Classification System to be a
significant advancement in tunneling as it introduced two new ideas, one of an active
unsupported span and second the concept of stand-up time, both of which are very helpful
to determine the type of support required in tunnels.
2.9.3. The Rock Quality Designation (Deere et al., 1967)
In 1967 Deere et al., developed an index called Rock Quality Designation for quantitative
estimation of rock mass quality from borehole logs of drilled cores. The Rock Quality
Designation (RQD) is defined as the percentage of rock pieces which are intact and have
length more than 100 mm found in the total length of core. It is a measure of the
fracturing degree (Dyke, 2006). The requirements in the Rock Quality Designation index
method includes that the double-tube core barrel should be used while drilling and
diameter of the core should not be less than 54.7 mm (NX-size). In present practice, the
RQD is a parameter of standard geotechnical core logging and provides a quick and cheap
index value of rock quality especially in weak rocks. In short, it is a measurement of the
percentage of good quality rock. Figure 2.9 shows the procedure to calculate the RQD
from a core.
CHAPTER-2 LITERATURE REVIEW
27
Figure 2.9: Calculation of RQD (After Deere, 1989)
In calculating the RQD index, only intact core is considered which has broken along the
boundaries of discontinuities. Drill breaks and breaks as a result of handling of the drill
cores which are actually not natural, are ignored. This avoids an underestimation of the
RQD index and consequently, of the quality of rock mass. The relationship between
quality of a rock mass and RQD index by Deere is presented in Table 2.2. The RQD
index is used in many geotechnical applications as well as in Q system of rock mass
classification. However, main drawback of the RQD index is that a high RQD value may
not always be an indication of high quality rock (Milne et al., 1989). Therefore the RQD
represents partially the quality of rock mass.
Although, RQD has been widely accepted as a measure of degree of fracture of the rock
mass, however it is a directionally dependent parameter and its value may change notably,
depending upon the orientation of the borehole.
CHAPTER-2 LITERATURE REVIEW
28
Table 2.2: The relationship between RQD and rock mass quality
Rock Quality Designation (%) Rock Mass Quality
<25 Very Poor
25 – 50 Poor
50 – 75 Fair
75 – 90 Good
90 – 100 Excellent
Palmström (1982) made suggestion that, when the core is not available but only the
discontinuity traces are noticeable in exploratory Adits or in surface exposures, the RQD
may be estimated indirectly by counting the number of discontinuities in a unit volume.
The suggested relationship for rock masses free of clay is:
𝑅𝑄𝐷 = 115 − 3.3𝐽𝑣 (2.5)
Where 𝐽𝑣 is the total of the number of joints in a unit length.
2.9.4. The Rock Structure Rating (Wickham et al., 1972)
Wickham used the case histories of relatively small tunnels supported by steel nets to
develop this classification system (Milne et al., 1998). Rock Structure Rating (RSR)
system introduced the idea of parameters based on rating to produce a numerical value for
rock quality (Dyke, 2006). Most of the case histories, used in the development of this
system, were for relatively small tunnels supported by means of steel sets, however, RSR
system was the first having reference of shotcrete support.
The Rock Structure Rating (RSR) is defined by the equation:
𝑅𝑆𝑅 = 𝐴 + 𝐵 + 𝐶 (2.6)
Where:
A = the geology parameter
B = the geometry parameter
CHAPTER-2 LITERATURE REVIEW
29
C = the effect of groundwater inflow along with joint condition
To calculate the resultant RSR value out of a maximum of 100, the parameter rating
values are assessed using tables which have been developed by Wickham et al., (1972),.
The tables are presented as Tables 2.2, 2.3 and 2.4.
Table 2.3: Rock Structure Rating - Parameter A
Basic Rock Type
Geological Structure Hard Medium Soft Decomposed
Igneous 1 2 3 4
Massive
Slightly
Folded
or
Faulted
Moderately
Folded or
Faulted
Intensively
Folded or
Faulted
Metamorphic 1 2 3 4
Sedimentary 2 3 4 4
Type 1 30 22 15 9
Type 2 27 20 13 8
Type 3 24 18 12 7
Type 4 19 15 10 6
Table 2.4: Rock Structure Rating - Parameter B
Average Joint
Spacing
Strike Perpendicular to Dip Strike Parallel to Axis
Direction of Derive Direction of Derive
Both With
Dip Against Dip Either Direction
Flat Dipping Vertical Dipping Vertical Flat Dipping Vertical
1. Very closely jointed
< 2 in 9 11 13 10 12 9 9 7
2. Closely jointed 2-6
in 13 16 19 15 17 14 14 11
3. Moderately joined
6-12 m 23 24 28 19 22 23 23 19
4. Moderate to blocky
1-2 ft 30 32 36 25 28 30 28 24
5. Blocky to massive
2-4 ft 36 38 40 33 35 36 24 28
6. Massive > 4 ft 40 43 45 37 40 40 38 34
(a) Dip flat 0’-20’, dipping 20’-50’ and vertical 50’-90’
CHAPTER-2 LITERATURE REVIEW
30
Table 2.5: Rock Structure Rating - Parameter C
Anticipated Water Inflow
gpm/1000 ft of Tunnel
Sum of Parameters
A+B
13 – 44 45 - 75
Joint Condition (b)
Good Fair Poor Good Fair Poor
None 22 18 12 25 22 18
Slight < 200 gpm 19 15 9 23 19 14
Moderate 200-1000 gpm 15 22 7 21 16 12
Heavy > 1000 gpm 10 8 6 18 14 10 (b) Joint Condition: good = tight or cemented; fair = slightly weathered or altered; poor = severely weathered , altered or open
In spite of some limitations, it is worth to note that the RSR system reveals the reason
behind developing a quantitative rock mass classification system. Even though the RSR
classification system is not extensively used today, the work of Wickham et al., had a
major role in development of some other classification systems being used in the world
today (Dyke, 2006).
2.9.5. Geomechanics or Rock Mass Rating System (Bieniawski, 1973)
The Geomechanics, or Rock Mass Rating (RMR) system was originally developed at the
South African Council of Scientific and Industrial Research (CSIR) (Singh et al., 1999).
The system was based on experience gained from shallow tunnel projects in sedimentary
rocks. The system was modified in 1976 and 1989. Bieniawski was of the view that no
classification systems which had been developed up to 1973, fully satisfied basic
requirements of a robust classification system. He pointed out two main limitations of the
classification systems available at that time. Many of the classification systems were
based wholly on the rock mass characteristics, and as such, were not practical. Also those
classification systems which were considered as practical did not include information
about rock mass properties. These systems therefore, were recommended by him to only
be applied to a single type of rock structure.
Like the most of classification systems of rock mass before it, Rock Mass Rating (RMR)
system was developed initially for use in civil engineering projects of underground
excavation and tunneling. The RMR System was an effort to have an extensive
classification system, which can fulfill the majority of practical requirements. This was
CHAPTER-2 LITERATURE REVIEW
31
done by combining the best features from the available respective classification systems
(Dyke, 2006).
For deciding the parameters to be used in a rock mass classification system of a jointed
rock mass, Bieniawski (1973) had the conclusion that since the design of engineering
structures in rocks requires prior site exploration, the required geotechnical parameters for
the classification of a rock mass should be obtained from data of a site investigation.
Consequently, the RMR system proposed by Bieniawski (1973) incorporated the
following six parameters:
i. Uniaxial compressive strength
ii. Rock Quality Designation (RQD)
iii. Spacing of discontinuities
iv. Discontinuities condition
v. Groundwater conditions.
vi. Discontinuities Orientation.
To apply RMR classification system the rock mass is grouped into several structural areas
and each area is classified individually. The boundaries of these areas are usually decided
by a major structural feature such as a fault or with a change in rock type (Hoek, 2005). In
some cases the procedure to divide the rock mass into many small structural portions can
facilitate the classification procedure.
According to Bieniawski, 1989, the most critical condition should be considered in case
of non-uniform conditions. In case two or more clearly distinct zones are present at small
scale through a unit to be considered homogeneous, then the overall weighted value based
on the area of each zone in relation to the whole area should be considered.
The detail of Rock Mass Rating system is presented in Table 2.5 and 2.6 showing the
ratings for each of the above listed parameters. All the six ratings are summed to give a
value of RMR as described in Table 2.7 whereas, Table 2.8 illustrates the meaning of the
five (5) rock mass classes giving ranges for stand up time, cohesion and friction angle of
the rock mass.
CHAPTER-2 LITERATURE REVIEW
32
Table 2.6: Input parameters of RMR (After Bieniawski, 1989)
PARAMETER Range of values // RATINGS
1 Strength
of intact
rock
material
Point-load
strength
index
> 10 MPa 4 – 10
MPa
2 – 4
MPa
1 – 2
MPa
For this low range
uniaxial compr.
Strength is preferred
Uniaxial
compressive
strength
> 225
MPa
100-250
MPa
50-100
MPa
25-50
MPa
5-25
MPa
1 – 5
MPa
< 1
MPa
RATING 15 12 7 4 2 1 0
2. Drill core quality RQD 90-100% 75-90% 50-75% 25-50% <25%
RATING 20 17 13 8 5
3. Spacing of discontinuities > 2m 0.6 – 2m 200–600
mm
60-200
mm
< 60 mm
RATING 20 15 10 8 5
4 Condition
of
discontinu
ities
Strength
Persistence
< 1 m 1-3 m 3-10 m 10-20 m > 20 m
Rating 6 4 2 1 0
Separation none < 0.1 mm 0.1 – 1
mm
1 -5 mm > 5 mm
Rating 6 5 4 1 0
Roughness Very
rough
rough Slightly
rough
smooth Slickensided
Rating 6 5 3 1 1
Infilling
(gouge)
none Hard Filling Soft Filling
- < 5 mm > 5 mm < 5 mm > 5 mm
Rating 6 4 2 2 0
Weathering Un-
weathered
Slightly
w.
Moderate
ly w.
Highly
w.
decomposed
Rating 6 5 3 1 0
5 Ground
water
Inflow per
10 m tunnel
length
none < 10
litre/min
10-25
Ltr./min
25-125
Ltr./min
>125 Ltr./min
pw/σ1 0 0-0.1 0.1-0.2 0.2-0.5 > 0.5
General
conditions
Comp. dry damp wet dripping flowing
RATING 16 10 7 4 0
pw = joint water pressure: σ1 = major principal stress
Table 2.7: Rating adjustment for discontinuity orientations
Very
favourable
Favourable Fair Unfavourable Very
unfavourable
RATINGS
Tunnels 0 -2 -5 -10 -12
Foundations 0 -2 -7 -15 -25
Slopes 0 -5 -25 -50 -60
Table 2.8: Rock mass classes determined from total ratings
Rating 100-81 80-61 60-41 10-21 < 20
Class No. I II III IV V
Description Very good Good Fair Poor Very poor
CHAPTER-2 LITERATURE REVIEW
33
Table 2.9: Meaning of rock mass classes
Class No. I II III IV V
Average stand-up time 10 years for
15 m span
6 months for
8 m span
1 week for
5 m span
10 hours for
2.5 m span
30 minutes
for 1 m span
Cohesion of the rock mass >400 kPa 300-400 kPa 200-300
kPa
100-200 kPa <100 kPa
Friction angle of the rock
mass
< 45° 35-45° 25-35° 15-25° <15°
The original RMR system proposed by Bieniawski in 1973 has been refined subsequently
and slight changes have been made several times during 1974 to 1989.
Many authors have modified the basic RMR System for specific applications (Dyke,
2006), including:
• Mining applications: Laubscher (1977, 1993) and Kendorski et al., (1983).
• Coal mining: Ghose and Raju (1981), Newman (1981), Unal (1983), Venkateswarlu
(1986) and Sheorey (1993).
• Slope stability: Romana (1985).
• The RMR value was related to the original Hoek-Brown equation in the making of
the Hoek-Brown failure criterion (Hoek and Brown, 1980).
The major advantage of the RMR system is its simplicity in use, while the main
disadvantages of the system are:
• In very poor rock masses, the system has been found undependable (Singh and Goel,
1999).
• The classification system is not sensitive to small variations in quality of the rock
mass.
• The classification system has been considered as being too conservative by the
mining engineers.
2.9.6. Rock Quality Index (Barton et al., 1974)
Rock Quality Index, also known as Q-System was derived in the Norwegian Geotechnical
Institute (NGI) which was based on over 200 tunnel case histories and underground
caverns (Singh and Goel, 1999). The system was developed by Barton, Lunde and Lien
(1974) for the design of support requirements for tunnels. In Q-System, six parameters are
CHAPTER-2 LITERATURE REVIEW
34
used to determine the quality of a rock mass. The rock quality index (Q) is calculated
from the equation:
Q =RQD
Jn .
JrJa .
JwSRF (2.7)
Where:
RQD is the Rock Quality Designation.
Jn = Joint Set number (number of discontinuities).
Jr = Joint Roughness number (roughness of the most unfavourable discontinuity).
Ja = Joint Alteration number (degree of the alteration or filling along the weakest
discontinuity).
Jw = Joint Water Reduction factor (water inflow into excavation).
SRF = Stress Reduction Factor (in situ stress condition).
The Q-System does not directly consider the rock mass strength. However this factor is
taken into account in Stress Reduction Factor (SRF). The SRF is derived from the
equation:
SRF = UCSσ′ (2.8)
Where:
UCS is the Uniaxial Compressive Strength
σʹ is the major principal stress
The Q index value can be described by three proportions, as shown in the followings:
• RQD
Jn
• Jr
Ja
• Jw
SRF
The first factor RQD/Jn corresponds to the rock mass structure, and is a rough measure of
the block size (Barton et al., 1974). The second factor Jr/Ja corresponds to the frictional
and roughness characteristics of walls of joint or the gouge materials. It is a crude
CHAPTER-2 LITERATURE REVIEW
35
indication of the inter block shear strength. The third part of the equation Jw/SRF is a
complex empirical factor having two stress parameters, and is an indication of the active
stress conditions. The SRF represents the total stress parameter and is a measure of:
• The release load in excavations through shear zones.
• Rock stress in good quality rock.
• Squeezing loads in case of weak plastic rock masses.
In the equation of Q index, water pressure is represented by the parameter Jw, which has a
negative effect on the joints shear strength by reducing the effective normal stress. The Q-
System does not consider the allowance for joint orientation which is a major exclusion.
Barton et al., (1974) considered that joint orientation is of less importance as was
expected initially. The input parameters of Q system are elaborated in the following
tables.
Table 2.10: Rock Quality Designation
Rock Mass Quality RQD (%)
Very Poor 0 – 25
Poor 25 – 50
Fair 50 – 75
Good 75 – 90
Excellent 90 – 100
Notes:
i. Where RQD is reported or measured as < 10 (including 0), a
nominal value of 10 is used to evaluate Q
ii. RQD intervals of 5 are sufficiently accurate
CHAPTER-2 LITERATURE REVIEW
36
Table 2.11: Joint set number (Jn)
Joint Set Jn
Massive, no or few joints 0.5 – 1
One joint set 2
One joint set plus random 3
Two joint sets 4
Two joint sets plus random 6
Three joint sets 9
Three joint sets plus random 12
Four or more joint sets, heavily jointed, “sugar-cube”, etc. 15
Crushed rock, earthlike 20
Notes: (i) For tunnel intersections, use (3.0xJn), (ii) For portals, use (2.0xJn)
Table 2.12: Joint roughness number (Jr)
a. Rock-wall contact,
b. Rock-wall contact before 10 cm shear
Discontinuous Joints Jr = 4
Rough or irregular, undulating 3
Smooth, undulating 2
Slickensided, undulating 1.5
Rough or irregular, planar 1.5
Smooth, planar 1.0
Slickensided, planar 0.5
Notes: i. Description refers to small scale features, and intermediate scale features, in
that order
c. No rock-wall contact when sheared
Zone containing clay minerals thick enough to prevent rock-wall contact Jr = 1
Sandy, gravelly or crushed zone thick enough to prevent rock-wall
contact
1
Notes:
i. Add 1.0 if the mean spacing of the relevant joint set is greater than 3 m.
ii. Jr = 0.5 can be used for planar slickensided joints having lineations, provided the
lineations are orientated for minimum strength
CHAPTER-2 LITERATURE REVIEW
37
Table 2.13: Joint alteration number (Ja) C
on
trac
t b
etw
een
jo
int
wal
ls JOINT WALL CHARACTER Condition Wall contact
CLEAN
JOINTS
Healed or welded
joints:
Filling of quartz, epidote, etc. Ja = 0.75
Fresh joint walls: No coating or filling, except from
staining (rust)
1
Slightly altered joint
walls:
Non-softening mineral coatings, clay-
free particles, etc.
2
COATING
OR THIN
FILLING
Friction materials: Sand, silt calcite, etc. (non-softening) 3
Cohesive materials: Clay, chlorite, talc, etc. (softening) 4
Par
tly
or
no w
all
con
tact
FILLING OF: Type Partly wall
contact
No wall
contact
Thin filling
(<5 mm)
Thick filling
Friction Materials Sand, silt, calcite, etc. (non-softening) Ja = 4 Ja = 8
Hard cohesive
materials
Compacted filling of clay, chlorite,
talc, etc.
6 5 – 10
Soft cohesive materials Medium to low overconsolidated
clay, chlorite, talc, etc.
8 12
Swelling clay materials Filling material exhibits swelling
properties
8-12 13-20
Table 2.14: Joint water reduction factor (Jw)
Description and ratings for the parameter Jw (joint water reduction factor)
Dry excavations or minor inflow, i.e. < 5l/min locally pw < 1 kg/cm2 Jw = 1
Medium inflow or pressure, occasional outwash of joint fillings 1 – 2.5 0.66
Large inflow or high pressure in competent rock with unfilled joints 2.5 - 10 0.5
Large inflow or high pressure, considerable outwash of joint fillings 2.5 - 10 0.3
Exceptionally high inflow or water pressure at blasting, decaying with
time
> 10 0.2 – 0.1
Exceptionally high inflow or water pressure continuing without
noticeable decay
> 10 0.1 – 0.05
Note: (i) The last four factors are crude estimates. Increase Jw if drainage measures are installed
(ii) Special problems caused by ice formation are not considered.
Table 2.15: Stress reduction factor (SRF)
A.
Wea
kn
ess
zo
nes
in
ters
ecti
ng
exca
vat
ion
Multiple weakness zones with clay or chemically disintegrated rock, very
loose surrounding rock (any depth)
SRF = 10
Single weakness zones containing clay or chemically disintegrated rock depth
of excavation < 50 m)
5
Single weakness zones containing clay or chemically disintegrated rock (depth
of excavation > 50 m)
2.5
Multiple shear zones in competent rock (clay-free), loose surrounding rock
(any depth)
7.5
Single shear zones in competent rock (clay-free), loose surrounding rock
(depth of excavation < 50 m)
5
Single shear zones in competent rock (clay-free), loose surrounding rock
(depth of excavation > 50 m)
2.5
Loose, open joints, heavily jointed or “sugar0cube” , etc. (any depth) 5
Continued...
CHAPTER-2 LITERATURE REVIEW
38
Note: (i) Reduce these value of SRF by 25 – 50% if the relevant
shear zones only influence, but do not intersect the
excavation
B
. C
om
pet
ent
rock
, ro
ck
stre
ss p
rob
lem
s
Low stress, near surface, open joints >200 0.01 2.5
Medium stress, favourable stress condition 200-10 0.01-
0.3
1
High stress, very tight structure. Usually favourable to
stability, may be except for walls
10-5 0.3-
0.4
0.5-2
Moderate slabbing after > 1 hour in massive rock 5 - 3 0.5 –
0.65
5 - 50
Slabbing and rock burst after a few minutes in massive rock 3 – 2 0.65 -
1
50 – 200
Heavy rock burst (strain burst) and immediate dynamic
deformation in massive rock
< 2 > 1 200 –
400
(i)
Note:
(ii)
For strongly anisotropic stress field (if measured): when 5 <σ2 /
When
Few case records available where depth of crown below surface is less
than span width. Suggest SRF increase from 2.5 to 5 for low stress
cases.
C. Squeezing rock Plastic flow of incompetent
rock under the influence of
high pressure
Mild squeezing rock
pressure
1 -5 5 – 10
Heavy squeezing rock
pressure
> 5 10 – 20
D. Swelling rock Chemical swelling activity
depending on presence of
water
Mild swelling rock
pressure
5 – 10
Heavy swelling rock
pressure
10 - 15
The Q index value varies from 0.001 to 1.000 on a logarithmic scale. The rock mass
quality is divided into nine classes. A summary of all the nine classes is shown in Table
2.16.
Table 2.16: Summary of Q-system classification (After Barton et al., 1990)
Q Index Value Rock Mass Class
0.0001 – 0.01 Exceptionally Poor
0.01 – 0.1 Extremely Poor
0.1 – 1.0 Very Poor
1.0 – 4.0 Poor
4 – 10 Fair
10 – 40 Good
40 – 100 Very Good
100 – 400 Extremely Good
400 – 1000 Exceptionally Good
CHAPTER-2 LITERATURE REVIEW
39
Apart from a modification in 1994 to the parameter SRF and the modifications made in
2002, the original parameters of the classification system remain unchanged (Singh and
Goel, 1999). According to Milne et al., (1998), the advantages of the Q system are:
• It is sensitive to minor changes in rock mass properties.
• The descriptors are thorough with little room for subjectivity.
The primary limitations of the Q system are:
• Inexperienced users can experience difficulty with the Jn parameter in a rock mass.
For example in widely jointed rock masses, in which an overestimation of the
number of joint sets in a rock mass can result in an underestimation of the Q index
(Milne et al., 1998).
• The SRF parameter is regarded as the most debatable parameter. Kaiser et al., (1986)
are of the opinion that in the rock mass classification, the SRF should not be
included, with the harmful effects of high stress being evaluated separately (Singh
and Goel, 1999).
2.9.7. The Geological Strength Index (Hoek et al., 1995)
The GSI system is a simple illustration method of classifying a rock mass for different
geological conditions. The system is composed of a chart with a description of a range of
rock mass structures. A sketch of the representative structure is on the vertical axis and on
the horizontal axis there is a description of a range of joint surface conditions. The GSI
value is determined from the relationship of an appropriate rock mass description and
joint surface description for a specific rock. The System GSI was introduced by Hoek
(1994), Hoek et al., (1995) and Hoek & Brown (1998) to complement the generalised
Hoek-Brown rock failure criterion. And to estimate the parameters s, a and mb used in the
criterion (Edelbro, 2003). The major advantage of the GSI system is that it helps a quick
classification of a rock mass.
In early days, the value of GSI was estimated directly from RMR. However, this
correlation has proved to be unreliable, particularly for poor quality rock masses. The
general GSI chart is presented as 2.10.
CHAPTER-2 LITERATURE REVIEW
40
Figure 2.10: The Geological Strength Index chart (Cai et al., 2004)
CHAPTER-2 LITERATURE REVIEW
41
According to Cai et al., (2004), the GSI system is the only rock mass classification
system that is directly linked to some engineering parameters like Mohr-Coulomb, Hoek-
Brown and rock mass modulus. However, due to its subjective nature, the application of
the GSI system is limited.
2.9.8. The Rock Mass Index (RMi) (Palmström 1995)
The rock mass index RMi, was initially proposed by Palmström in 1995. After that it has
been further modified and refined. RMi is a parameter which is volumetric and this
parameter indicates the approximate uniaxial compressive strength of a rock mass. The
value of RMi is applied as input parameter for estimating rock support requirement and to
other rock engineering problems (Palmström, 2009)
There is some similarity in input parameters between RMi and Q system. For example,
the joint and its features are almost the same. The input parameters for RMi can be
obtained by field observations and measurements. Major disadvantage of the system is
that it requires more calculations than the Q system and RMR.
RMi uses of the uniaxial compressive strength of intact rock (σc) and the effect which
reduces the joints penetrating the rock mass (JP), is given by:
RMi = σc . JP (2.9)
Where:
σc = uniaxial compressive strength of the intact rock,
JP = Jointing parameter determined by empirical relations JC (joint conditions) and Vb
(block volume). Charts are available to evaluate the RMi values for the rock mass.
As the RMi value characterizes the strength properties of the dry rock mass material, it
does not take into account the influence from stresses of the rock and ground water.
(Palmström, 2009)
The RMi system is best for jointed, massive and crushed rock mass where the joints in the
various sets are of similar properties. The system may also be used in over-stressed and
brittle ground (Palmström, 1995). Some limitations have also been identified by
Palmström like great care in the categorization and estimate of rock support in complex
CHAPTER-2 LITERATURE REVIEW
42
and weak zones. RMi should be applied with care in special zones like squeezing ground
while the swelling ground is not dealt with in this system.
2.9.9. Rock Mass Number and Rock Condition Rating
As a part of the development of the existing rock mass classification systems, two new
parameters have been developed. One has been adopted from Q system and is called as
Rock Mass Number (N) and other one from Rock Mass Rating and is known as Rock
Condition Rating (RCR). N can be defined as follows:
𝑁 =RQD
Jn .
JrJa . Jw (2.10)
So if we eliminate SRF from Q system equation, N can be derived. This was required as
there were many uncertainties while calculating the SRF in Q system.
Similarly RCR is defined as the RMR of a rock mass without rating of the crushing
strength of the intact material and some alteration of the orientation of joints (Singh and
Goel 1999).
RCR = RMR – (crushing strength rating + adjustment of the orientation of joints)
Parameter wise both N and RCR are comparable and can be used for inter- relation.
2.9.10. Slope Mass Rating
The Slope Mass Rating (SMR) was proposed from new geomechanical classification i.e.
RMR for rock slopes (Romana, 1985). The classification is obtained from the RMR-
system by using an adjustment factor which depends upon the relation between the slope
and joints. Another factor was used depending on the method of excavation. Like other
classification systems, the SMR determines the need of support and explains the rock in
five different classes.
2.10. COMPARISON OF CLASSIFICATION SYSTEMS
As different classification systems give importance to different parameters, it is therefore
recommended that at least two systems should be used when classifying a rock mass
CHAPTER-2 LITERATURE REVIEW
43
(Hoek, 2000). The parameters included in the four of the classification systems which
have been used in this research, are compared in Table 2.17 (Edelbro, 2003).
Table 2.17: Parameters included in different classification systems
Parameter Classification Systems
RMR Q RSR GSI
No. of Joint Set
Joint Spacing
Joint Strength
Rock Type
State of Stress
Groundwater Condition
Strength of the intact rock
Blast Damage
2.11. CORRELATIONS BETWEEN ROCK CLASSIFICATIONS SYSTEMS
Many researchers have worked on the classification systems of the rock mass and
suggested several correlations. Most of the researchers have worked on correlations of
RMR and Q system, both being the most famous systems. However some other systems
like RSR and GSI have also been inter-related to each other by some researchers.
2.11.1. Significance of Correlations
The main rock mass classification systems utilize the similar rock mass parameters to
some extent. Consequently, it is possible to compare these systems but with some
limitations. For example the estimated rock support for an underground project
determined in a system can be checked and compared in the other systems. Such
CHAPTER-2 LITERATURE REVIEW
44
comparisons lead to more reliable estimates, provided the characterization of the ground
is carefully carried out (Milne, 1998).
For rock engineering and design, Bieniawski (1989) suggests to apply at least two
classification systems when determining the empirical tools.
Due to the common usage of classification systems, a number of statistical correlations
have been developed by many researchers to relate the rock mass rating values derived
from different systems to one another. Usually, rock mass classification data are not
always available in a form that may immediately be applied to a specific engineering
problem. Therefore, correlations may be very useful to rapidly derive different design
aspects. Furthermore, the availability of correlation equations between classification
systems facilitates a rapid means of verifying resultant rock mass rating values, without
re-calculation of the values (Dyke, 2006).
2.11.2. Correlation between RMR and Q System
RMR and the Q system are the main classification systems for estimates of rock support.
Both systems use the most vital ground features for input parameters. Every parameter is
classified individually and each rating expresses the quality of the rock with respect to
stability of underground structure (Palmström, 2009).
Several empirical correlations between RMR and Q system have been developed based
on case histories in different countries. The first effort was made by Bieniawski in 1976
to correlate RMR with Q values. He analysed over one hundred case histories (68 in
Scandinavian countries, 21 in USA and 28 in South Africa). Although this relation has
been widely used in practice, several other relations were suggested in the following
years. This depended on the fact that, such kind of relations are site sensitive and
therefore cannot be generalised. It should be noted that those correlations are only based
on a statistical basis and their physical bases are different. Care should be taken when
applying these relations for dissimilar rock conditions.
Some of the relations between RMR and Q system collected from the literature are listed
in Table 2.18 below:
CHAPTER-2 LITERATURE REVIEW
45
Table 2.18: Correlations between RMR and Q system (Kennert Röshoff et al., 2002,
Dyke, 2006)
Correlation Developed / Referred By:
𝑅𝑀𝑅 = 9𝑙𝑛𝑄 + 44 (2.11) Bieniawski, 1976
𝑅𝑀𝑅 = 5.9𝑙𝑛𝑄 + 43 (2.12) Rutledge and Preston, 1978
𝑅𝑀𝑅 = 5.4𝑙𝑛𝑄 + 55.2 (2.13) Moreno, 1980
𝑅𝑀𝑅 = 5𝑙𝑛𝑄 + 60.8 (2.14) Cameron-Clarke and Budavari, 1981
𝑅𝑀𝑅 = 10.5𝑙𝑛𝑄 + 41 (2.15) Abad, 1984
𝑅𝑀𝑅 = 13.5𝑙𝑛𝑄 + 43 (2.16) Milne et al., 1989
𝑅𝑀𝑅 = 12.5𝑙𝑜𝑔𝑄 + 55.2 (2.17) Milne et al., 1989
𝑅𝑀𝑅 = 43.89 − 9.9𝑙𝑛𝑄 (2.18) Milne et al., 1989
𝑅𝑀𝑅 = 12.11𝑙𝑜𝑔𝑄 + 50.81 (2.19) Milne et al., 1989
𝑅𝑀𝑅 = 8.7𝑙𝑛𝑄 + 38 (2.20) Milne et al., 1989
𝑅𝑀𝑅 = 10𝑙𝑛𝑄 + 39 (2.21) Milne et al., 1989
𝑅𝑀𝑅 = 15𝑙𝑜𝑔𝑄 + 50 (2.22) Barton, 1995
𝑅𝑀𝑅 = 7𝑙𝑛𝑄 + 36 (2.23) Tugrul, 1998
𝑅𝑀𝑅 = 9𝑙𝑛𝑄 + 49 (2.24) Al-Harthi, 1993
Some other correlations suggested by different researchers are as follows;
2.11.3. Correlation between RSR and Q System
Rutledge & Preston (1978) worked on many case histories of New Zealand and developed
following correlation between RSR and Q system.
𝑅𝑆𝑅 = 13.3𝑙𝑛𝑄 + 46.5 (Rutledge & Preston, 1978) (2.25)
CHAPTER-2 LITERATURE REVIEW
46
Turgul (1998) studied the clayey limestone on which the Ataturk dam has been founded
in Turkey. He divided the rock mass into three classes and suggested the following
correlation.
𝑅𝑆𝑅 = 6𝑙𝑛𝑄 + 46 (Tugrul, 1998) (2.26)
𝑅𝑆𝑅 = (4𝑙𝑛𝑄 + 51) ± 8 (Jauch, 2000) (2.27)
2.11.4. Correlation between RSR and RMR
Following correlations have been found between RSR and RMR in the literature.
𝑅𝑆𝑅 = 0.77𝑅𝑀𝑅 + 12.40 (Rutledge & Preston, 1978) (2.28)
𝑅𝑆𝑅 = 0.78𝑅𝑀𝑅 + 17 (Tugrul, 1998) (2.29)
𝑅𝑀𝑅 = (0.7𝑅𝑆𝑅 + 29) ± 5 (Jauch, 2000) (2.30)
2.11.5. Correlation between GSI and RMR
GSI is a relatively new system, therefore less literature has been found. However
following correlations of GSI have been found with RMR
𝐺𝑆𝐼 = 0.69𝑅𝑀𝑅 + 4.71 (Milne et al., 1989) (2.31)
𝐺𝑆𝐼 = 𝑅𝑀𝑅 − 5 (Hoek et al., 1995) (2.32)
2.12. CORRELATIONS BETWEEN ROCK CLASSIFICATIONS SYSTEMS AND
MODULUS OF DEFORMATION
As the in situ tests to determine the deformation modulus are costly, time consuming and
require special procedures, there have been some attempts to correlate the modulus with
rock mass classification system. Bieniawski in 1978 made the first empirical model for
prediction of the modulus of deformation of rock mass. After Bieniawski’s equation,
some researchers developed other empirical approaches with other systems like RSR, GSI
and Q system. A summary based on the rigorous literature search is shown in Table 2.19.
CHAPTER-2 LITERATURE REVIEW
47
Table 2.19: Correlations between Modulus of Deformation and different rock mass
classification systems
Sr. No Correlation No. Developed By:
Correlations with RMR
1. Em = 2RMR – 100 when RMR > 50 (2.33) Bieniawski, 1978
2. Em = 10(RMR - 10)/40
when RMR ≤ 50 (2.34) Serafim & Pereira, 1983
3. Em = Ei/100 (0.0028RMR
2 + 0.9
exp(RMR/22.82)) (2.35)
Nicholson & Bieniawski,
1990
4. Em = Ei (0.5 (1 – cos (π.RMR/100))) (2.36) Mitri & Edrissi, 1994
5. Em = 0.1 (RMR/10)3
(2.37) Read et al., 1999
6. Em = (1- D/2) 𝜎𝑐𝑖
100 × 10
𝑅𝑀𝑅 −10
40 (2.38) Hoek et al., 2002
7. Em = 0.0003RMR
3 – 0.0193RMR
2 +
0.315RMR + 3.4064 (2.39)
Muhammadi &
Rehmannejad, 2010
8. Em = Ei e(RMR-100)/36
(2.40) Bieniawski, 2007 &
Galera et al., 2007
9. Em/Ei = 1/100 (0.0028RMR
2 +
0.9RMR/22.82
) (2.41)
Nicholson & Bieniawski,
1990
10. Em = 19.43 ln(RMR) – 69.03 (2.42) Kayabasi et al., 2003
11. Em = 0.0736e0.0755RMR
(2.43) Gokcoeoglu, 2003
Correlations with Q System
12. Em = 25log10 Q for Q>1 (2.44) Barton, 1993
13. Em = 10 𝑄𝑐
1
3 ; Qc =UCS/100 (2.45) Barton, 2000
14. Em = 8 Q0.4
for 1<Q<30 (2.46) Palmstrom, 2001
15. Em = 15log10 Q + 50 (2.47) Barton, 2002
Correlations with RSR
16. Log10 Em = 10(RSR - 52)/109
(2.48) Sarma, 2005
Correlations with GSI
17. Em = 1.989 ln(D) – 2.512
D = Ei (1-RQD×GSI) (2.49) Ghamgosar, 2010
18. Em = 0.912 e0.866GSI
(2.50) Gokcoeoglu, 2003
CHAPTER-2 LITERATURE REVIEW
48
19. Em = 0.145 e0.0645GSI
(2.51) Ghamgosar, 2010
20. Em = Ei (0.02 + 1−
𝐷
2
1+ 𝑒 75+25𝐷−𝐺𝑆𝐼
11
) (2.52) Hoek & Diedrich, 2006
21. Em = 𝜎𝑐′
100 10
𝐺𝑆𝐼 −10
40 (2.53) Hoek et al., 1998
22. Em = 0.804 e0.0386GSI
(2.54) Kayabasi et al., 2003
23. Em = Ei (S)
1/4 ; Ei = 50GPa ; S =
exp(GSI – 100/9) (2.55) Carvalho, 2004
24. Em = tan ( 1.56 + ln𝐺𝑆𝐼 2) 𝜎𝑐𝑖3
(2.56) Beiki et al., 2010
25.
Em = Ei (Sa)0.4
S = exp(GSI - 100)/a
a = 0.5 + 1/6(e-GSI/15
– e-20/3
)
(2.57) Sonmez, 2004
Correlations with Modulus of Elasticity/RQD
26. Em = αE Ei
αE = 0.0231RQD – 1.32 ; (≥0.15) (2.58) Gardner, 1987
2.13. SUMMARY
It is essential to determine the physical and mechanical properties of rocks for an
engineering project. The modulus of deformation is one of the parameters which represent
the mechanical behaviour of a rock mass. In fact this parameter is considered to be more
important than strength of the rock mass. Plate loading and plate jacking tests are mostly
recommended by the experts to determine the modulus of deformation in the field.
There are about twelve classification systems of rock mass which were developed
between 1946 and 2002 and which are being used successfully for characterization and
design of underground excavations. Most of the researchers have worked on RMR and Q
system to correlate the two schemes with each other. Fewer correlations have been found
among other systems being less used.
All these expressions shown in Tables 2.18 and 2.19 have arisen from a series of specific
data taken from some limited data base. Therefore use of these correlations with extreme
caution about the compatibility of the data has been recommended by many researchers.
CHAPTER-3
49
ROCK PROPERTIES OF THE STUDY AREAS
3.1. INTRODUCTION
Knowledge about the physical and mechanical properties of the rock mass is of great
importance for reducing the potential problems and disturbance during construction of the
structures over the rocks or within the rocks. This would help in better understanding of the
failure process and a better rock mass strength and deformation predictions (Edelbro,
2003). The properties are used in characterizing and classification of the rock mass. Hoek
(2007) has described that strength of a jointed rock mass depends upon the properties of the
intact rock pieces and also on the freedom of these pieces to rotate and slide under different
stresses. So reliability of the strength and deformation characteristics greatly depends upon
true identification of the rock mass properties.
In this research geological and geotechnical investigations of Diamer Basha dam and
Kohala Hydropower Project sites have been analysed. Both these projects are located in the
northern area of Pakistan having different types of rocks. Basha dam site consists of
intrusive igneous rocks while Kohala site has sedimentary rocks.
This chapter presents the review of the geological and geotechnical studies carried out at
both the sites, rock types, detail of laboratory tests and properties of rock mass using
RocLab software.
3.2. ROCK PROPERTIES OF DIAMER BASHA DAM SITE
Diamer Basha Dam has been planned at the Indus River, between the Khyber Pakhtunkhwa
Province and the Northern Areas, approximately 315 km upstream of Tarbela Dam, about
165 km downstream of the Northern Area capital Gilgit and some 40 km downstream of
Chilas (Figure 3.1). The Project consists of a 270 m high Roller Compacted Concrete
(RCC) dam and two Hydroelectric Power Schemes at either side of the Indus River. Both
power schemes comprise an extensive and complex network of underground works
including power cavern, transformer and switchgear cavern, headrace and tailrace tunnels,
surge tanks, access and diversion tunnels (DBC, 2007).
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
50
Figure 3.1: Location plan of Diamer Basha Dam and Kohala Hydropower Project sites
3.2.1. General Geology
The Diamer Basha Dam project is situated within the Jurassic–Cretaceous island arc in
northern Pakistan known as Kohistan Arc. The rock types exposed in the reservoir have
undergone extensive deformation due to the high degree of tectonic activity of the region
(Monenco, 1998). The prevailing rock type at the site is a mafic intrusive rock which is
petrologically called Diorite or Gabbronorite (GN). In the field the rock appears very strong
and massive. The fresh hand specimen is comparably heavy which is proven by laboratory
testing, revealing an average density of 2.89 g/cm3. The fresh rock is rather light coloured.
Usually the GN is grey to light grey, but it is varying due to changes in quantitative mineral
compositions. In some areas the rock is significantly darker coloured than usual. A rusty
layer is covering the rock in some areas which is particularly different from the desert
varnish. Among the minerals that can be identified with field methods the plagioclases,
pyroxenes and amphiboles are dominating (DBC, 2007). Figure 3.2 shows a close up view
of a specimen which is typical for GN.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
51
Figure 3.2: A close view of typical Gabbronorite rock piece
Another rock formation is also present at the site which is called Ultramafic Association
(UMA) having mafic minerals more than 90%. The rock types grouped under this term
reveal also a very diverse nature (DBC, 2008). They are even heavier having density of
3.23 g/cm3, which can be felt in the hand specimen. Their strength is also high but not
reaching that of the GN. The rock is more intensively weathered, but it is hardly weakened.
Somewhere, UMA has an intensive rusty colour because of weathering of Iron-bearing
minerals. The joints are often stained with calcite. The UMA rocks seem to be part of the
main injections of mantle derived magmas in the Chilas Complex GN rocks. A close up
view of typical UMA sample obtained from the site is shown in Figure 3.3.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
52
Figure 3.3: A close view of a UMA rock sample
3.2.2. Geotechnical Investigations at Basha Dam Site
The area of the dam has been investigated by several means of exploration. The main
information had been gathered by drilling and water pressure testing. More than 16000 m
core drilling in 120 boreholes was carried out. A borehole scanner system (ETIBS®) had
been used in 29 boreholes within the dam footprint or close to it. Six trenches (total length
331.5 m) have been excavated to collect information about colluvial soils. Five (5) Adits
with total length of more than two kilometres have been excavated. Geological mapping of
the Adits has been done along with rock sampling.
Among the five Adits, two exploratory Adits have been driven on either side of the Indus
River with total lengths of 532 m for left bank Adit (Adit 4) and 651 m for right bank Adit
(Adit 5) to investigate the area of the envisaged power caverns. The Adits consist of a main
drive and of cross-cuts, perpendicular to the main drive. Both Adits have a standard cross
section with a width of 2.4 m, a height of 3.2 m having circular crown. The initial 150 m of
Adit 4 runs in south-west direction. Thereafter the Adit turns northwest to follow the axis
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
53
of the main cavern. The cross cut starts at chainage 316.55 m. From the portal of the Adit 4,
the rock mass is massive but has a very complicated joint pattern. The joint spacing is in
the range of 1 to 3 m. Most of the joints are tight and show no infill. The spacing between
open joints with infill is 6 to 7 m. The infilling consists mainly of weathered pegmatites and
silty fillings with minor clay amounts. Little seepage can be observed along the some joint
planes.
The first 125 m of Adit 5 run in north-east direction. Subsequently the Adit turns into the
direction south-east (N120°E) to follow the axis of the main cavern. The conditions in the
access part of Adit 5 are favourable with massive GN. The rock is jointed though and some
of the discontinuities have a persistence of greater than 10 m. The Adit intersects three
steeply inclined fracture zones, which might be evidence for stress relief. These zones can
possibly be connected with small depressions and erosional gullies at surface (DBC, 2008).
Locations of the Adit 4 and Adit 5 are represented in Figure 3.4 on layout plan of the
project while Figures 3.5, 3.6 and 3.7 show the photographs of Adit 4 and core examination
at the Basha site respectively.
Figure 3.4: Layout plan of Diamer Basha Dam showing Adit 4 and 5
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
54
Figure 3.5: Portal and inside view of the Adit 4 of Diamer Basha Dam
Figure 3.6: Core examination for Diamer Basha Dam
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
55
Figure 3.7: Core examination and selection for laboratory testing for Diamer Basha Dam
3.2.3. Laboratory Testing
For assessment of the engineering geological properties of the rock and to obtain
parameters for the geotechnical design, a number of different rock mechanic tests have
been carried out. The cores were carefully selected and preserved as per standard
procedures. Three major campaigns of laboratory tests have been performed at Central
Material Testing Laboratories (CMTL) WAPDA Lahore. In addition, shear box tests and
point load strength index tests were carried out at site.
Index Tests
Index testing included the standard evaluation methods for determining unit weight,
specific gravity, water absorption and porosity. The tests have been carried out based on the
recommendations given by ISRM. Index tests have been conducted on a total of 77
samples. Figure 3.8 shows the number of Index tests performed.
UMA Gabbronorite
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
56
Figure 3.8: Index property tests for Diamer Basha Dam
All the index tests have been summarized and the mean values with ranges are presented in
Table 3.1.
Table 3.1: Summary of the index properties
Rock
Type
Specific Gravity Unit Weight
(KN/m3)
Water
Absorption (%) Porosity
Range Mean Range Mean Range Mean Range Mean
Gabbro
-norite 2.84-3.45 2.94 27.4-34.1 28.6
0.046-
1.030 0.22 0.14-3.35 0.65
UMA 2.82-3.54 3.29 27.9-34.8 31.7 0.066-
1.830 0.70 0.20-10.00 2.22
Unconfined Compression Strength, Young’s Modulus and Poisson Ratio
A total of 106 tests were performed for the determination of UCS. Testing procedure was
based on the ISRM – Suggested Methods for Determining the Uniaxial Compressive
0
20
40
60
80
100
120
Unit Weight Water Absorption Specific Gravity Porosity Water Content
No
. o
f T
ests
Index Property Tests
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
57
Strength and Deformability of Rock Materials. Both types of the major hard rock
lithologies have been tested. Of the total tested samples, 35 belong to the UMA, while 71
samples belong to the Gabbronorite rocks.
The unconfined compressive strength of the specimen is obtained by dividing the
maximum load carried by the specimen during test by the cross-sectional area and the result
is reported to the nearest 10 psi (68.9 kPa).
For determining the deformability of the core pieces, 79 tests have been performed with
strain gauges fixed on the core specimen. The deformation has been measured
continuously. The applied standard is the same as mentioned above. These tests have also
been evaluated in order to obtain the Young’s modulus (Ei) and Poisson’s ratio (𝜐) of the
rock. The value of Poisson’s ratio is greatly affected by the non-linearities at low stress
level in the axial and lateral stress strain curves.
Point Load Strength Index Testing
Point Load Strength Index Tests (PLSIT) were carried out on all the 106 samples submitted
for laboratory testing (total 106). These were performed for getting a reliable conversion
factor between UCS and PLSIT and based on the ISRM – Suggested Methods for
Determining Point Load Strength. Additionally a point load machine was deployed at the
site. During its operation 434 samples of different boreholes have been tested.
Tensile Strength
The applied procedure was based on the guidelines from ISRM – Suggested Methods for
Determining Tensile Strength of Rock Materials. Ten samples were chosen for that
purpose. All the samples were taken from boreholes in the riverbed. The test values are also
termed as splitting tensile strength. The test specimen had length-to-diameter L/D ratio of
½, cut from a drilled core. The length of the specimen should be at least 10 times greater
than the largest mineral grain constituent. Care was taken that the thickness of the disk
should be greater than the largest mineral grain constituent.
Typical test result sheets of engineering properties tests carried out in CMTL Lahore, are
placed in Appendix A. Figure Nos. 3.9 to 3.13 represent the different laboratory testing
performed on samples from Diamer Basha Dam.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
58
Figure 3.9: Preparation of samples by cutting the cores
Figure 3.10: Point Load Strength Index test
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
59
Figure 3.11: Preparations for Modulus of Elasticity test
Figure 3.12: Unconfined Compression Test without strain gauges
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
60
Figure 3.13: Indirect tensile strength test
All the tests were performed by following ASTM or ISRM standards. The summary
showing total number of each engineering property test performed is shown in Figure 3.14.
Figure 3.14: Engineering properties tests performed on cores of Basha site
0
20
40
60
80
100
120
Point Load
Strength Index
Test
Uniaxial
Compressive
Strength
Modulus of
Elasticity
Poisson’s Ratio Tensile Strength
No. of
Tes
ts
Engineering Properties
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
61
The results of the engineering properties tests performed on selected cores are presented in
Appendix B.
For the rocks of Basha, relatively high values for UCS were expected; therefore the load
intervals were spaced such that about 20 readings were taken per sample in order to get
smooth stress-strain curves. Load intervals were evenly spaced and mentioned in the test
results. The tests had cycles of loading and unloading before finally increasing the stress
until failure of the cores. In some of the tests, stress fluctuations were seen near the peak
strength.
The averages values of each test have been chosen as the representative. The data for the
GN shows much less scatter than that for the UMA, which is probably due to the high
diversity of UMA and their varying contents and kinds of mafic minerals. The mean values
selected on the basis of laboratory tests are given in Table 3.2.
Table 3.2: Engineering properties of intact rock material for Diamer Basha Dam
Rock
Type
Unconfined
Compressive
Strength (MPa)
Young’s
Modulus (GPa)
Poisson Ratio PLSIT (MPa)
Range Mean Range Mean Range Mean Range Mean
Gabbro
-norite 29-203 100 3.7-250 60
0.017-
0.952 0.25 1.71-14 5.2
UMA 15-138 80 13.2-340 100 0.022-
0.887 0.26 1.29-12.6 4.8
3.2.4. Properties of Rock Mass Using RocLab Software
To determine the rock mass strength parameters, results from the laboratory tests were
extrapolated to the rock mass with the help of RocLab® software. The input data consists
of;
unconfined compressive strength of intact rock, sigci (UCS)
the intact rock parameter mi
the geological strength index GSI
the disturbance factor D
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
62
These parameters accommodate damages due to blasting or any other excavation method.
For description of the rock mass the GSI (Geological Strength Index) is used. The value
can be picked by a chart (Figure 2.9, Chapter 2) which describes the structural appearance
and conditions of the rock mass. Input data such as D, mi, UCS and Ei were kept constant
for each rock types and for a range of different GSI values, parameters like global (rock
mass) strength, rock mass uniaxial compressive strength, rock mass tensile strength and
modulus of deformation have been calculated from RocLab. The results are shown in Table
3.3.
Table 3.3: Summary of results of rock mass strength for Basha Dam using RocLab
Rock Type GSI
Rock Mass
Tensile
Strength
(MPa)
Rock Mass
UCS
(MPa)
Modulus of
Deformation
(GPa)
Global
Strength
(MPa)
Gabbronorite
D=0.1, mi=23,
UCS=100 MPa,
Ei=60 GPa
40 0.042 2.94 6.89 19.99
50 0.091 5.47 13.35 24.85
60 0.197 9.91 23.13 30.74
70 0.427 17.75 33.50 38.43
UMA
D=0.1, mi=25,
UCS=80 MPa,
Ei=100 GPa
40 0.031 2.35 13.79 16.68
50 0.067 4.37 26.71 20.71
60 0.0145 7.93 46.27 25.59
70 0.314 14.20 66.99 31.91
The Disturbance Factor, D has been selected as 0.1 for the application in “Tunnels” which
is for good qulity controlled blasting resulting in some disturbance to the confined rock
mass surrounding the tunnel. The values of mi for both the rocks have been picked from the
charts available in the software while the uniaxial compressive strength and intact modulus
have been taken from the mean values of the rocks based on the lab testing described in
Table 3.2. The range of values of GSI has been used from 40 to 70 i.e. for a good to very
good quality rock found at Diamer Basha site. The results show that as the GSI rating
values increase, the other parameters also increase.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
63
A typical plot from RocLab having input / output parameters is shown in Figure 3.15. The
plot also shows the two curves showing the two relations; first between major and minor
principal stress and second between normal and shear stress.
Figure 3.15: A typical plot from RocLab
If required, cohesion and friction angle of the rock mass can also be determined along with
other paramters from RocLab. The relation between GSI and global strength has been
plotted in Figure 3.16.
The Figure shows that UMA has better global strength than Gabbronorite for same GSI
ratings. Similar relations can be obtained in the form of graphs from the results shown in
Table 3.4.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
64
Figure 3.16: The relation between GSI and global strength for rocks of Basha
3.3. ROCK PROPERTIES OF KOHALA HYDROPOWER PROJECT SITE
Kohala Hydropower Project area lies in Muzaffarabad district of AJK on the River Jhelum
(Figure 3.1). A large syntaxeal bend in AJK territory, known as Domel bend with its apex
at Muzaffarabad has been used to generate the power through 16.6 km long tunnel. The
dam site is located on the upper limb at Siran, and power house on the lower limb of river
Jhelum at Barsala. (Fig.3.17). The selected layout of the Kohala Hydropower project
comprises 52 meters high concrete gravity dam with crest length of 160 meters at 904.5
meters above sea level. The area between Siran Dam site and Agar Nullah near Barsala is
characterised by high mountains with peaks up to 2100 meter height, however the tunnel
will have its maximum cover up to the elevation of 1990 meter. The dam intake area is at
the elevation of about 850 meters and Agar Nullah is flowing at an elevation of 980 meters.
Muzaffarabad-Kohala area is tectonically very active. This region is characterized by a
number of regional faults including the Panjal Thrust, the Main Boundary Thrust, the
Muzaffarabad Thrust and the Jhelum Fault.
0
5
10
15
20
25
30
35
40
45
30 40 50 60 70 80
Glo
bal
Str
ength
(M
Pa)
GSI
Gabbronorite
UMA
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
65
Figure 3.17: Layout plan of Kohala Hydropower Project
3.3.1. General Geology
The rocks exposed in the area are sandstone and shale which belong to Murree formation.
Sandstone has been classified into two types i.e. Sandstone 1 (SS-1) and Sandstone 2 (SS-
2) striking from NW-SE, dipping from 50º to more than 80º towards NE. The shale is
mostly interbedded with SS-2. It is reddish brown in colour, fine grained and comparatively
less hard. These rocks are sedimentary which are secondary in their origin, the materials of
which they are composed having been derived from the decay and disintegration of some
previously existing rock mass (KHC, 2009). The characteristics of rock units are as
follows;
Sandstone-1 (SS-1)
This is the dominant rock unit in the Powerhouse area and along the tunnel route. This rock
is not very well exposed in the intake area. The rock is generally fresh, fine to medium
grained, well cemented and hard. As seen in the core samples and field expressions, joints
are mostly tight, with few joints having filling and coating of calcite.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
66
Sandstone-2 (SS-2)
This sandstone is a transitional unit between grey colour SS-1 and Shale. The fine grained
rock is generally reddish brown in colour but sometimes its colour gradually changes from
grey to dull grey and then to reddish brown. It is medium hard and comparatively thin
bedded. The rock looks highly weathered on surfaces, but in core samples it is generally
fresh to slightly weathered. It has silty clayey contents at places mostly near the contact
with shale.
Shale
The Shale present at site is mostly interbedded with SS-2. It is reddish brown in colour, fine
grained and comparatively less hard. At places it is splintery with well developed
laminations.
3.3.2. Geotechnical Investigations at Kohala Hydropower Project
Intensive field investigations have been carried out during the feasibility stage of the
project. The geological and geotechnical investigations were accomplished through
geological mapping, core drilling at different sites of the project, geophysical survey,
excavation of two Adits and in situ testing in the Adits. Laboratory tests on rock cores,
aggregate and river sand samples have been carried out to evaluate the geotechnical
parameters for the design and construction of the project. The investigation mostly
concentrated to key areas for the project, such as;
The main dam
Desander chambers
Diversion tunnel and the intake facilities
The headrace tunnel crossing with Agar Nullah
Power station area at Barsala.
Two exploratory Adits, one located at the left abutment of the main dam, and the other at
access tunnel at powerhouse site have been excavated. The length of both Adits is 200 m
each with a maximum height of 3.2 m and width of 2.5 m.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
67
The Adit 1 is excavated on the left bank of Jhelum River to investigate the area of the left
abutment and desanding chambers. The encountered rock is mainly SS-1, SS-2 and Shale.
The Adit 2 is excavated near to the access / surge tunnel in the vicinity of the Barsala
powerhouse site. This Adit is exactly at the level of the access tunnel, which will be later on
used during tunnel construction. This Adit is designed especially to investigate the headrace
tunneling conditions in the area. The encountered rock is mainly SS-1, SS-2 and Shale. In
Adit 2, also the rock mechanic in situ tests were carried out.
The rock cores obtained from drilling in different area were preserved by following the
standard procedures. Figures 3.18 and 3.19 show the examination of cores for Kohala HPP.
Figure 3.18: Core examination for Kohala Hydropower Project
Characterisation of the rock cores of selected borehole is presented in Table 3.4. In most of
the boreholes, degree of jointing is high to very high resulting in RQD < 50%. Only BH-9
and BH-26 have predominantly a moderate degree of jointing which correspond to a range
of RQD from 50% to 75%.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
68
Figure 3.19: Core examination and selection for laboratory testing for Kohala HPP
Table 3.4: Rock mass characteristics of selected boreholes of Kohala HPP.
Borehole No.:
Degree of jointing/RQD (%)
Very high High Moderate Low Very low
RQD<25 25-50 50-75 75-90 >90
BH-01 44 40 10 4 2
BH-02 62 29 3 6 0
BH-03 36 28 26 7 3
BH-04 37 50 13 0 0
BH-05 94 6 0 0 0
BH-08 100 0 0 0 0
BH-09 5 32.5 40 10 12.5
BH-10 100 0 0 0 0
BH-11 27.5 40 27.5 2.5 2.5
BH-12 42.5 35 20 2.5 0
BH-13 30 20 10 30 10
BH-15 34 43 23 0 0
BH-26 4 38.5 38.5 16.5 2.5
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
69
3.3.3. Laboratory Testing
For assessment of the engineering geological properties of the rock and to obtain
parameters for the geotechnical design, several rock mechanics tests have been carried out.
The representative core samples were selected for laboratory testing in order to determine
the index properties and other engineering characteristics. Rock mechanics testing has been
performed on representative cores samples selected from each major lithological unit to
characterize the range of properties.
The tests were performed in Central Material Testing Laboratories WAPDA Lahore on
selected rock core samples. A comprehensive laboratory testing program was carried out
for this purpose. For selecting the rock core samples from the boreholes, the boreholes were
divided in various zones along the borehole depth. For example, an over burden zone, stress
support zone, structural / excavation zone and the foundation zone. Sufficient number of
laboratory tests from each zone were conducted in order to characterize each lithological
unit and to evaluate the engineering characteristics of the important zones as described
above.
Figure 3.20 represents the number of each test performed on selected core samples. The
tests have been carried out based on the recommendations given by ISRM.
Figure 3.20: Tests performed on core samples of Kohala HPP
0
20
40
60
80
100
120
140
Index
Properties
Point Load
Strength Index
Uniaxial
Compressive
Strength
Modulus of
Elasticity
Poisson’s
Ratio
Tensile
Strength
No. of
Tes
ts
Index and Engineering Properties Tests
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
70
Summary of the all the index tests along with mean values of the three rock units are
presented in Table 3.5.
Table 3.5: Summary of the index properties of Kohala site – mean values
Rock Type Specific
Gravity
Unit Weight
(KN/m3)
Water
Absorption
(%)
Porosity
SS-1 2.74 27 0.50 0.04
SS-2 2.68 25 2.30 0.05
Shale 2.71 24 3.50 0.06
Mean specific gravity, unit weight and porosity are similar in all three types of rocks. The
percentage water absorption is in different range having low value for SS-1 and high for
SS-2 and Shale.
The results of the engineering properties tests performed on selected cores are presented in
Appendix B. Based on the laboratory test results, mean representative values have been
selected which are shown in Table 3.6.
Table 3.6: Engineering properties of intact rock material of Kohala – mean values
Rock
Type
Unconfined
Compressive
Strength (MPa)
Young’s
Modulus
(GPa)
Poisson
Ratio
PLSIT
(MPa)
SS-1 80 40 0.20 8
SS-2 50 30 0.15 5
Shale 20 25 0.10 3
3.3.4. Properties of Rock Mass of Kohala using RocLab Software
Like Basha, global strength and modulus of deformation have been calculated for Kohala
by using input data such as compressive strength, D, mi and Ei for each rock unit and for a
range of different GSI values. Input data such as compressive strength, D, mi and Ei were
kept constant for each rock types and for a range of different GSI values, parameters like
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
71
global strength, rock mass uniaxial compressive strength, rock mass tensile strength and
modulus of deformation have been calculated from RocLab. The results are based on a
range of GSI from 20 to 50 i.e. for a poor to good quality rock mostly found at Kohala site.
Because of the frequency and number of encountered joint sets, the basic framework of the
rock mass is described as “blocky” grading into “very blocky” in weaker parts. The surface
conditions were observed to be good as most of the joints are only slightly weathered. The
results are shown in Table 3.7 which describes that as the GSI rating values increase, the
strength parameters also increase for all three types of rocks.
Table 3.7: Summary of results of rock mass strength for Kohala site using RocLab
Rock Type GSI
Rock Mass
Tensile
Strength
(MPa)
Rock Mass
UCS (MPa)
Modulus of
Deformation
(GPa)
Global
Strength
(MPa)
Sandstone-1
D=0, mi=17,
UCS=80 MPa,
Ei=40 GPa
20 0.011 0.64 1.83 8.66
30 0.024 1.38 3.25 11.54
40 0.051 2.65 6.39 14.56
50 0.108 4.82 12.29 18.00
Sandstone-2
D=0, mi=17,
UCS=50 MPa,
Ei=30 GPa
20 0.007 0.40 1.37 5.41
30 0.015 0.80 2.44 7.21
40 0.032 1.65 4.79 9.10
50 0.068 3.01 9.22 11.25
Shale
D=0, mi=6,
UCS=20 MPa,
Ei=25 GPa
20 0.008 0.16 1.14 1.23
30 0.017 0.34 2.03 1.69
40 0.036 0.66 3.99 2.17
50 0.077 1.21 7.68 2.75
The relation between GSI and global strength has been plotted in Figure 3.21.
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
72
Figure 3.21: GSI vs Global Strength for Kohala HPP
The Figure shows that shale is the weakest in global strength for same GSI rating values,
while SS-1 is stronger in two types of sandstones.
3.4. SUMMARY
Diamer Basha dam and Kohala hydropower project are proposed in northern area of
Pakistan. Extensive geological and geotechnical investigations have been carried out at
both the sites. The laboratory testing has been supervised by the author in order to get the
representative rock mechanics parameters. The index and engineering properties for both
the sites have been summarized to get the mean values. Based on the geological data of
Basha and Kohala hydropower project sites, it can be inferred that Basha dam site mainly
comprises two types of rocks mass namely Gabbronorite (GN) and Ultramafic Association
(UMA). The high values of RQD for these rocks (75 to 85%) indicate that the Basha dam
site rocks are classified from good to very good. At Kohala hydropower project site, three
types of rock units exist, i.e., Sandstone-1 (SS-1), Sandstone-2 (SS-2) and Shale. The RQD
values for Kohala rocks ranges between 20 – 45% indicating poor to fair quality rocks.
0
2
4
6
8
10
12
14
16
18
20
10 15 20 25 30 35 40 45 50 55
Glo
bal
Str
ength
(M
Pa)
GSI
SS-1
SS-2
Shale
CHAPTER-3 ROCK PROPERTIES OF THE STUDY AREAS
73
The results from the laboratory tests were extrapolated to the rock mass with the help of
RocLab software and values for the generalized Hoek Brown Criterion were computed.
Rock mass parameters were generated including tensile strength, uniaxial compressive
strength, rock mass strength and deformation modulus for both the projects. The relations
between GSI and global strength for both the sites have also been plotted. The results show
that Gabbronorite has better global strength for same GSI ratings for Basha site while shale
is the weakest in global strength for same GSI values at Kohala site. Also Sandstone-1 is
stronger in two types of sandstones at Kohala.
CHAPTER-4
74
CORRELATIONS BETWEEN VARIOUS ROCK MASS
CLASSIFICATION SYSTEMS
4.1. INTRODUCTION
The major classification systems like RMR and Q system use the most important ground
features as input parameters such as RQD, condition and spacing of the discontinuities
and groundwater etc. Each of these parameters is classified individually and each rating
expresses the quality of the rock. However there is also much dissimilarity between these
two systems and also among all the classification systems.
The data sets regarding rock mass are not always available in a format that may
immediately be applied to a specific engineering problem. Therefore correlations between
different systems may be very useful to rapidly derive different design parameters.
Furthermore, the availability of correlation equations between classification systems
facilitates a rapid means of verifying resultant rock mass rating values, without requiring
the re-calculation of the values. For example rock support found in one system can be
checked in other system. In this chapter classification systems have been applied to both
sites and useful correlations have been developed.
4.2. CLASSIFICATION SYSTEMS APPLIED IN THE STUDY
In this research, the rock masses of Diamer Basha Dam and Kohala Hydropower Project
sites have been classified by four main and well known rock mass classification systems
i.e. RMR, Q System, RSR and GSI. Detail of each system, parameters involved and
methodologies for the respective determinations have been discussed in Chapter 2.
4.3. ROCK MASS CLASSIFICATION OF DIAMER BASHA DAM SITE
For Diamer Basha Dam, laboratory testing data and geological mapping of both the Adits
(Adit 4 and 5) have been used as the input parameters to classify the rock mass. The total
combined length of both the Adits is 1138 m. The rock classification has been done at an
interval of 25 meter along the length of both the Adits as per ASTM Designation D5878-
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
75
08 (2008). A sample portion of mapping is shown in Fig. 4.1 whereas, geological
mapping from Ch: 75 to 225 is shown in Appendix C.
Figure 4.1: Typical geological mapping (Ch: 138 – 150) of Adit 4 of Basha site
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
76
4.3.1. Parametric Study of the Rocks of Basha Site
The required strength parameters for the classification systems can be determined by
laboratory testing, or by geological judgement through field methods. The orientations of
the discontinuities can either be assessed by surface joint surveys, exploratory Adits or
borehole logging and scanner surveys. A combination of these methods is also applicable
and removes bias from the gathered data. The influence of the joint orientation with
respect to the tunnel is accommodated by applying a range from unfavourable to
favourable orientations to the RMR value. The ground water conditions are more difficult
to judge. In cases, where the rock mass classification is based solely on drilling this factor
is rather difficult to estimate.
To classify the rock mass in RMR system, average uniaxial compressive strength
(unconfined compressive strength) determined from the laboratory tests (Table 3.2) has
been used as strength of the intact rock material. The RQD gives directly the rating values
of each portion of the Adit. For Basha, average RQD values of 85 for Adit 4 and 75 for
Adit 5 have been used. The joint spacing in Adit 4 with no infilling is in the range of 1 to
3 m. The spacing between open joints with infill is 6 to 7 m. The infilling consists mainly
of weathered pegmatites and silty fillings with minor clay amounts. The condition in Adit
5 is also similar. However, some of the discontinuities have a persistence of greater than
10 m. Minor seepage or wet spots can be observed along some joint planes in both the
Adits. The ratings of these parameters have been carefully done accordingly keeping in
view all the factors stated above.
For classification in Q system, same RQD values have been used as in RMR. Joint set
number has been selected for a massive rock having few joints, while joint roughness
number has been taken for rough, irregular and undulating surface. Most of the joints in
both the Adits have thin coating of non softening material like silt etc. The joints in the
main GN are sometimes displaying intensively weathered fillings in both the Adits, but
their thickness is very limited at surface. As there are minor inflows in both the Adits,
joint water reduction factor has been selected as close to 1.0 mostly. Similarly since the
area has favourable stress conditions, stress reduction factors have been chosen as 1.0 or
close to 1.0. The influence of the stress relief on the low angled joint sets can be
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
77
demonstrated on the example of the discontinuity systems encountered in both the
exploratory Adits.
Rock Structure Rating (RSR) system having less parameters is relatively easy to apply.
For slightly faulted to massive igneous rock at Basha, first parameter has been selected
accordingly. For second parameter blocky to massive rock has been selected with dip of
the prominent joints between 20o to 50
o. Third parameter describes the anticipated water
inflow which is dry to slight and joint condition tight to slightly weathered. So the rating
values have been picked accordingly.
The values of GSI rating for each particular portion are directly derived from the chart
given in Figure 2.9 (Chapter 2). The surface condition of the joints is mostly rough and
slightly weathered and the structure is blocky in both the Adits.
Worksheets have been prepared for each of the four systems used to calculate the
numerical value for each portion being classified. All the required parameters have been
incorporated to calculate the numerical values. Accordingly the rock quality of each
portion has been determined in all the systems.
The high compressive strength of the Gabbronorite, which is the most abundant rock at
site, together with the lack of groundwater are more or less generally valid for most of the
rock mass in both Adits. The “good rock” class further requires tight or closed
discontinuities with moderate to wide spacing which is persistent at the site. The joints are
healed and are mainly opened up due to blasting and drilling. Occasionally the material
within the joint is slightly weathered but this is not affecting the adjacent rock. The ”very
good class” is present in limited parts of the Adits 4 and 5. When a portion stretching over
a chainage of several metres is considered the rating usually drops to “good rock”. This is
because of the number of joints and joint sets encountered. As the difference between
very good and good rock is mainly based on the jointing, the general appearance of the
rock and the joint conditions are pretty similar between these two classes.
The methodologies to calculate the rating values of the rock mass in RMR, Q system
RSR and GSI have been discussed in chapter 2. The worksheets to determine ratings in
each of the four systems describing numerical values of each parameters and resultantly
quality of all portions of rock mass are shown in Table 4.1 to 4.4 in the following pages.
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
78
Table 4.1: Calculation of RMR values for Diamer Basha Dam site
Chainage
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Discontinuities
Condition of
Discontinuities Groundwater
Strike & Dip
Orientation of
Discontinuities
Main Adit 4
0-25 9 17 18 20 14 -2 76 Good
25-50 9 17 18 10 14 -5 63 Good
50-75 9 17 18 12 10 -5 61 Good
75-100 9 17 18 10 10 -5 59 Good
100-125 9 17 15 11 12 -5 59 Good
125-150 9 17 18 15 12 -5 66 Good
150-175 9 17 12 16 10 -2 62 Good
175-200 9 17 18 15 14 -5 68 Good
200-225 9 17 12 16 14 -5 63 Good
225-250 9 17 18 20 14 -5 73 Good
250-275 9 17 18 15 14 -5 68 Good
275-300 9 17 15 20 12 -5 68 Good
300-325 9 17 15 15 14 -5 65 Good
325-350 9 17 18 15 12 -5 66 Good
350-375 9 17 13 20 14 -10 63 Good
375-400 9 17 15 15 14 -5 65 Good
400-422 9 17 18 15 12 -5 66 Good
Right-X-Cut
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
79
Chainage
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Discontinuities
Condition of
Discontinuities Groundwater
Strike & Dip
Orientation of
Discontinuities
0-25 9 17 16 15 14 -8 63 Good
25-50 9 17 15 20 14 -8 67 Good
50-75 9 17 16 20 14 -5 71 Good
75-100 9 17 20 18 14 -5 73 Good
100-111 9 17 20 18 14 -5 73 Good
Main Adit 5
0-25 9 17 20 28 15 -2 87 Very Good
25-50 9 17 15 28 15 -2 82 Very Good
50-75 9 17 15 28 14 -2 81 Very Good
75-100 9 17 18 22 14 -5 75 Good
100-125 9 17 18 22 14 -7 73 Good
125-150 9 17 18 22 15 -5 76 Good
150-175 9 17 20 28 15 -2 87 Very Good
175-200 9 17 15 25 15 -5 76 Good
200-225 9 17 20 28 15 -5 84 Very Good
225-250 9 17 15 28 15 -5 79 Good
250-275 9 17 20 28 15 -3 86 Very Good
275-300 9 17 15 28 15 -5 79 Good
300-325 9 17 18 25 10 -5 74 Good
325-350 9 17 20 25 7 -3 75 Good
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
80
Chainage
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Discontinuities
Condition of
Discontinuities Groundwater
Strike & Dip
Orientation of
Discontinuities
350-375 9 17 18 22 15 -5 76 Good
375-400 9 17 20 23 15 -2 82 Very Good
400-425 9 17 15 23 15 -3 76 Good
425-451 9 17 18 20 15 -5 74 Good
Right-X-Cut
0-25 9 17 20 25 15 -3 83 Very Good
25-50 9 17 20 25 15 -5 81 Very Good
50-75 9 17 18 22 15 -3 78 Good
75-100 9 17 15 24 15 -3 77 Good
Left-X-Cut
0-25 9 17 18 20 15 -5 74 Good
25-50 9 17 20 23 15 -5 79 Good
50-75 9 17 18 22 15 -5 76 Good
75-100 9 17 18 20 10 -5 69 Good
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
81
Table 4.2: Calculation of Q index values for Diamer Basha Dam site
Chainage RQD Joint Set
No. (Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint Water
Reduction
Factor (Jw)
Stress
Reduction
Factor
(SRF)
Q Rock Quality
Main Adit 4
0-25 85 0.85 3.00 3.00 0.90 1.00 90.00 Very Good
25-50 85 0.85 3.00 3.00 0.85 1.50 56.67 Very Good
50-75 85 0.90 2.80 3.00 0.80 1.00 70.52 Very Good
75-100 85 0.90 3.00 2.80 0.66 1.00 66.79 Very Good
100-125 85 0.85 3.00 3.00 0.85 2.50 34.00 Good
125-150 85 0.90 3.00 3.00 0.80 1.00 75.56 Very Good
150-175 85 0.85 2.50 3.00 0.88 1.00 73.33 Very Good
175-200 85 0.80 2.50 2.80 0.90 1.00 85.38 Very Good
200-225 85 0.85 2.80 3.00 0.90 1.00 84.00 Very Good
225-250 85 0.90 3.00 3.00 0.80 1.00 75.56 Very Good
250-275 85 0.85 2.50 3.00 0.90 1.50 50.00 Very Good
275-300 85 0.88 2.50 3.00 0.90 1.50 48.30 Very Good
300-325 85 0.85 3.00 2.80 0.80 2.00 42.86 Very Good
325-350 85 0.90 3.00 3.00 0.90 1.00 85.00 Very Good
350-375 85 0.90 2.80 3.00 0.90 1.00 79.33 Very Good
375-400 85 0.80 3.00 2.50 0.90 2.50 45.90 Very Good
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
82
Chainage RQD Joint Set
No. (Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint Water
Reduction
Factor (Jw)
Stress
Reduction
Factor
(SRF)
Q Rock Quality
400-422 85 0.85 3.00 3.00 0.75 1.00 75.00 Very Good
Right-X-Cut
0-25 85 0.90 3.00 3.00 0.90 2.50 34.00 Good
25-50 85 0.90 3.00 3.00 0.90 2.50 34.00 Good
50-75 85 0.85 3.00 2.50 0.85 1.50 68.00 Very Good
75-100 85 0.85 3.00 3.00 0.90 1.00 90.00 Very Good
100-111 85 0.85 3.00 2.80 0.85 1.00 91.07 Very Good
Main Adit 5
0-25 75 0.85 3.00 2.20 0.85 1.00 102.27 Ext. Good
25-50 75 0.90 3.00 2.40 0.90 1.00 93.75 Very Good
50-75 75 0.85 3.00 2.31 0.80 1.00 91.67 Very Good
75-100 75 0.85 3.00 2.20 0.90 2.50 43.32 Very Good
100-125 75 0.85 3.00 2.00 0.88 1.50 77.65 Very Good
125-150 75 0.90 3.00 2.40 0.90 1.50 62.50 Very Good
150-175 75 0.80 2.65 2.00 0.95 1.00 118.01 Ext. Good
175-200 75 0.80 2.50 2.00 0.95 1.65 67.47 Very Good
200-225 75 0.80 3.00 2.00 0.90 1.00 126.56 Ext. Good
225-250 75 0.80 2.80 2.34 0.95 1.00 106.57 Ext. Good
250-275 75 0.80 2.30 2.00 0.91 1.00 98.11 Very Good
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
83
Chainage RQD Joint Set
No. (Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint Water
Reduction
Factor (Jw)
Stress
Reduction
Factor
(SRF)
Q Rock Quality
275-300 75 0.80 2.50 2.40 0.84 1.00 82.03 Very Good
300-325 75 0.85 3.00 2.20 0.80 1.00 96.26 Very Good
325-350 75 0.85 3.00 2.50 0.75 1.50 52.94 Very Good
350-375 75 0.85 3.00 2.50 0.95 1.00 100.59 Ext. Good
375-400 75 0.80 3.00 2.30 0.94 1.00 114.95 Ext. Good
400-425 75 0.88 3.00 2.50 0.85 1.00 86.93 Very Good
425-451 75 0.85 3.00 2.48 0.80 1.00 85.39 Very Good
Right-X-Cut
0-25 75 0.89 3.00 2.00 0.90 1.00 113.76 Ext. Good
25-50 75 0.80 2.50 2.50 0.91 1.00 85.31 Very Good
50-75 75 0.88 2.50 2.50 0.80 1.00 68.18 Very Good
75-100 75 0.85 3.00 2.50 0.95 2.50 40.24 Very Good
Left-X-Cut
0-25 75 0.85 3.00 2.50 0.85 1.00 90.00 Very Good
25-50 75 0.85 3.00 2.50 0.89 1.00 94.24 Very Good
50-75 75 0.85 3.00 2.50 0.95 1.00 100.59 Ext. Good
75-100 75 0.90 2.50 2.50 0.75 1.00 62.50 Very Good
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
84
Table 4.3: Calculation of RSR values for Diamer Basha Dam site
Chainage
Ratings for RSR =
A+B+C
Rock
Quality General Area
Geology (A)
Joint Pattern,
Direction of
Drive (B)
Groundwater,
Joint
Condition (C)
Main Adit 4
0-25 22 38 22 82 Very Good
25-50 20 38 22 80 Very Good
50-75 18 38 22 78 Good
75-100 18 35 20 73 Good
100-125 18 35 20 73 Good
125-150 20 35 20 75 Good
150-175 22 38 25 85 Very Good
175-200 20 35 23 78 Good
200-225 22 38 20 80 Very Good
225-250 20 35 21 76 Good
250-275 22 38 24 84 Very Good
275-300 20 35 22 77 Good
300-325 20 35 17 72 Good
325-350 20 35 17 72 Good
350-375 20 38 20 78 Good
375-400 22 38 20 80 Very Good
400-425 20 35 19 74 Good
425-451 22 38 21 81 Very Good
Right-X-Cut
0-25 15 32 19 66 Good
25-50 18 32 19 69 Good
50-75 20 32 20 72 Good
75-100 20 35 19 74 Good
100-111 18 35 19 72 Good
Main Adit 5
0-25 22 38 22 82 Very Good
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
85
Chainage
Ratings for RSR =
A+B+C
Rock
Quality General Area
Geology (A)
Joint Pattern,
Direction of
Drive (B)
Groundwater,
Joint
Condition (C)
25-50 20 38 22 80 Very Good
50-75 18 38 22 78 Good
75-100 18 35 20 73 Good
100-125 18 35 20 73 Good
125-150 20 35 20 75 Good
150-175 22 38 25 85 Very Good
175-200 20 35 23 78 Good
200-225 22 38 20 80 Very Good
225-250 20 35 21 76 Good
250-275 22 38 24 84 Very Good
275-300 20 35 22 77 Good
300-325 20 35 17 72 Good
325-350 20 35 17 72 Good
350-375 20 38 20 78 Good
375-400 22 38 20 80 Very Good
400-425 20 35 19 74 Good
425-451 22 38 21 81 Very Good
Right-X-Cut
0-25 20 38 22 80 Good
25-50 20 35 22 77 Good
50-75 20 35 20 75 Good
75-100 20 35 21 76 Good
Left-X-Cut
0-25 22 35 22 79 Good
25-50 20 38 22 80 Good
50-75 20 38 20 78 Good
75-100 22 35 19 76 Good
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
86
Table 4.4: GSI values for Diamer Basha Dam site
Chainage GSI Rock Quality
Main Adit 4
0-25 60 Good
25-50 46 Fair
50-75 45 Fair
75-100 47 Fair
100-125 48 Fair
125-150 52 Fair
150-175 48 Fair
175-200 52 Fair
200-225 50 Fair
225-250 53 Fair
250-275 50 Fair
275-300 48 Fair
300-325 56 Fair
325-350 52 Fair
350-375 50 Fair
375-400 56 Good
400-422 58 Good
Right X-Cut
0-25 47 Fair
25-50 48 Fair
50-75 58 Good
75-100 57 Good
100-111 56 Good
Main Adit 5
0-25 60 Good
25-50 64 Good
50-75 60 Good
75-100 62 Good
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
87
Chainage GSI Rock Quality
100-125 58 Good
125-150 52 Fair
150-175 58 Good
175-200 65 Good
200-225 53 Fair
225-250 64 Good
250-275 57 Good
275-300 62 Good
300-325 59 Good
325-350 56 Good
350-375 53 Fair
375-400 57 Good
400-425 62 Good
425-451 55 Fair
Right X-Cut
0-25 56 Good
25-50 63 Good
50-75 61 Good
75-100 62 Good
Left X-Cut
0-25 53 Fair
25-50 55 Fair
50-75 57 Good
75-100 57 Good
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
88
4.4. ROCK MASS CLASSIFICATION OF KOHALA HYDROPOWER
PROJECT SITE
For Kohala Hydropower Project site, data of fourteen (14) boreholes located at the site of
four (4) following different project structures has been used.
Main Dam BH No. 8,9,10,11,12
Desander BH No. 1,2,3,15
Diversion Tunnel BH No. 4,5,6,7
Powerhouse BH No. 26
Study of borehole logs and visual inspection of the cores was conducted for parametric
evaluation and rock mass classifications. Figure 4.2 shows typical core box while
borehole logs of BH No.11 and 12 of main dam area are presented in Figure 4.3. More
borehole logs are placed in Appendix D.
Figure 4.2: Core box of BH 15 (Depth 95 to 100 m) of Kohala site
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
89
Figure 4.3: Typical borehole logs of BH 11 and 12 showing lithology at Kohala site
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
90
4.4.1. Parametric Study of the Rocks of Kohala Site
Almost in all the boreholes, the discontinuity spacing is considered to be rather small due
to the low RQD values. Some of the boreholes e.g. BH-2, BH-9, BH-11 and BH-15 show
that a few portions are relatively wider spaced in terms of jointing; however there are
portions with denser spacing and shattered parts also.
The length of the discontinuities is difficult to judge from drilling. It is expected that
thicker portions with shattered rock are persistent for at least several meters, possibly
even in the range of tens of meters. On the other hand this might be true for a zone of
closer jointing or weakness, the picture for the individual joints making up such a zone
might be quite different. The joints are mostly planar, rough and weathered. Calcite is the
most common filling material. The heavily jointed rock masses are often related to weak
shales. It is assessed that most of the joints are altered and contain to a certain extent soft
infillings of calcite, silt or eventually even clay. Hard infillings of calcite or iron staining
can also be expected. Consequently it had been estimated that the separation between
such weathered joints is between 1 and 5 mm in this portion, giving a rating of 1.
The roughness of the surfaces is very much dependant on the thickness of the infilling
and its type. In order to accommodate for discontinuities with less filling and clean joints
without filling it was decided to go for a rating of 3 which corresponds to slightly rough
conditions. The infilling can be in the range of a few millimetres and hard as well as soft
infillings are expected, thus a rating of 2 gives a good compromise.
The difficulty when applying borehole data to the RMR and Q systems is that the RQD as
one of the main input parameters is strongly direction dependant. In case the borehole
was drilled parallel to any major joint or discontinuity, the RQD would be constantly
lower and pretend a much worse condition. Similarly, it has to be kept in mind that the
thickness of weak zones can be exaggerated if the bore is passing it at other angles than
perpendicular.
Large variations in most of the parameters of the RMR system have been observed as
most of the properties vary in large extent at various locations of Kohala project site. The
RQD values vary from 9 to 72. The spacing of discontinuities varies from less than 60
mm to more than 2 m. Condition of the discontinuities is rough and slightly to highly
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
91
weathered with infilling of soft material mostly. Groundwater rating also has large range
from completely dry to flowing under water table.
For Q system, joint set number parameter is selected for a “two joint set” to “three joint
set” rock. Joint roughness number varies for a smooth planar to rough or irregular rock-
wall contact. Infilling of clay, silt and even sand has been observed in some joints. The
rating for joint alteration number has been done accordingly. Joint water reduction factor
has been taken as 1.0 for dry condition to 0.66 for medium inflow. The variation in stress
reduction factor is even more. It varies from high stress level in some of the rock cores to
multiple weakness zones with clay or chemically disintegrated rock. The rating for the
adjustment of discontinuity orientations is set to fair as it is difficult to judge upon such
features when no orientation data is available.
RSR system parameters have also been evaluated carefully. General geology of the area
consists of slightly to moderately folded or faulted structure with occasionally occurrence
of intensively weathered rock. Mostly the strike is found perpendicular to the axis. The
joints are moderately to very closely spaced. The groundwater condition has been
selected from none (dry) to moderately flowing.
The values of GSI for each particular portion are directly derived from the chart given in
Figure 2.2 (Chapter 2). The joint surface condition found is mostly weathered or altered
and the structure is blocky / disintegrated in most of the cores. Therefore the values GSI
have been picked accordingly.
Similar worksheets as prepared for Basha site have been prepared for each of the four
systems to calculate the numerical values. All the required parameters have been
incorporated with due care to calculate the numerical values. Accordingly, the rock
quality of each portion has been determined in all the systems. The worksheets to
determine RMR, Q system and RSR for Kohala site are shown in Table 4.5 to 4.8.
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
92
Table 4.5: Calculation of RMR values for Kohala Hydropower Project site
BH #
Depth from
Ground
Surface (m)
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Disconti-
nuities
Condition
of Disconti-
nuities
Ground
-water
Strike & Dip
Orientation of
Discontinuities
Dam
8
11.7-20.7 4 3 5 1 10 -15 8 Very Poor
20.7-22.7 4 5 5 3 10 -7 20 Very Poor
22.7-34.7 4 3 5 7 10 -7 22 Poor
34.7-60.7 4 3 5 4 15 -15 16 Very Poor
9
15-26 7 5 8 2 15 -25 12 Very Poor
26-32 7 8 15 16 15 -25 36 Poor
33-34 7 8 15 20 15 -25 40 Poor
34-45.4 7 8 10 25 15 -25 40 Poor
45.4-49.4 7 5 10 8 15 -25 20 Poor
49.4-70 7 8 15 22 15 -25 42 Fair
70-71 7 8 20 20 15 -25 45 Fair
71-74 7 13 20 23 10 -25 48 Fair
74-75 12 17 20 20 15 -15 69 Fair
75-81 7 13 20 11 15 -25 41 Fair
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
93
BH #
Depth from
Ground
Surface (m)
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Disconti-
nuities
Condition
of Disconti-
nuities
Ground
-water
Strike & Dip
Orientation of
Discontinuities
81-84 12 13 20 24 15 -25 59 Fair
10
2.9-11.9 4 3 5 5 10 -2 25 Poor
11.9-18.7 4 5 8 7 10 -2 32 Poor
18.7-31 4 8 8 20 10 -2 48 Fair
31-40.9 2 5 5 5 10 -15 12 Very Poor
40.9-43.4 4 5 5 16 10 -2 38 Poor
43.4-51.9 7 5 8 21 10 -2 49 Fair
51.9-53.6 7 5 8 20 10 -2 48 Fair
53.6-61.9 4 5 5 19 10 -2 41 Fair
11
5-6 7 8 15 23 4 -25 32 Poor
6-7 12 13 15 27 4 -25 46 Fair
7-13 7 13 15 15 4 -25 29 Poor
13-20 2 5 5 21 0 -25 8 Very Poor
20-21 12 13 20 25 4 -15 59 Fair
21-25 4 5 5 21 4 -25 14 Very Poor
25-30 7 8 15 23 4 -25 32 Poor
30-34 7 13 15 26 4 -25 40 Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
94
BH #
Depth from
Ground
Surface (m)
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Disconti-
nuities
Condition
of Disconti-
nuities
Ground
-water
Strike & Dip
Orientation of
Discontinuities
34-35 12 13 20 28 4 -25 52 Fair
35-44 7 5 5 20 4 -25 16 Very Poor
44-50 12 13 20 25 4 -25 49 Fair
50-52 12 13 20 26 4 -15 60 Fair
52-56 7 8 20 25 4 -25 39 Poor
56-57 7 8 15 19 4 -25 28 Poor
57-63 12 8 20 22 4 -25 41 Fair
63-71 12 13 20 25 4 -25 49 Fair
71-84 7 13 20 21 4 -25 40 Fair
12
4-8 7 13 15 14 4 -25 28 Poor
8-16 7 8 10 13 4 -25 17 Very Poor
16-29 4 5 10 14 4 -25 12 Very Poor
29-36 7 5 10 16 4 -25 17 Very Poor
36-73 7 8 15 19 4 -25 28 Poor
Desander
1 45-51 7 8 13 26 10 -5 59 Fair
54-59 4 5 5 8 10 -5 27 Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
95
BH #
Depth from
Ground
Surface (m)
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Disconti-
nuities
Condition
of Disconti-
nuities
Ground
-water
Strike & Dip
Orientation of
Discontinuities
84-89 4 8 13 15 10 -5 45 Fair
104-114.4 7 8 13 18 10 -5 51 Fair
114.4-115.8 7 8 13 21 10 -5 54 Fair
115.8-116.8 7 8 13 24 10 -5 57 Fair
2
45-50 2 5 5 1 10 -10 13 Very Poor
53-58 2 3 5 1 10 -12 9 Very Poor
85-90 7 8 10 21 10 -5 51 Fair
105-110 7 8 8 14 10 -5 42 Fair
3
48-53 4 5 5 4 10 -10 18 Very Poor
53-58 4 5 5 4 10 -10 18 Very Poor
58-61 4 5 5 3 10 -10 17 Very Poor
85-86 7 5 10 15 10 -5 42 Fair
86-90 12 8 15 23 10 -5 63 Good
104-108.55 7 8 15 24 10 -5 59 Fair
108.55-111.6 7 13 15 22 10 -5 62 Good
111.6-115 7 13 15 20 10 -5 60 Fair
15 48-53 7 8 15 14 10 -2 52 Fair
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
96
BH #
Depth from
Ground
Surface (m)
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Disconti-
nuities
Condition
of Disconti-
nuities
Ground
-water
Strike & Dip
Orientation of
Discontinuities
56-61 7 8 15 15 10 -2 53 Fair
85-89 7 8 10 16 10 -2 49 Fair
89-90 7 8 10 17 10 -2 50 Fair
104-107 7 13 15 24 10 -2 67 Good
107-109 7 8 8 11 10 -2 42 Fair
109-110 7 8 10 16 10 -2 49 Fair
110-115 7 13 15 20 10 -2 63 Good
Diversion Tunnel
4
22.4-24.4 4 5 5 10 10 -2 32 Poor
24.4-25.4 4 5 5 11 10 -2 33 Poor
25.4-27.4 4 5 5 10 10 -2 32 Poor
37.4-39.1 4 5 5 10 10 -2 32 Poor
39.4-55.4 4 5 5 15 10 -2 37 Poor
5
55-56 2 5 5 3 10 -10 15 Very Poor
56-57 7 13 15 20 10 -5 60 Fair
57-58 2 5 5 4 10 -10 16 Very Poor
58-59.2 2 5 5 3 7 -10 12 Very Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
97
BH #
Depth from
Ground
Surface (m)
Ratings for
RMR Rock
Quality PLSI/
UCS RQD
Spacing of
Disconti-
nuities
Condition
of Disconti-
nuities
Ground
-water
Strike & Dip
Orientation of
Discontinuities
59.2-60.2 4 5 5 20 10 -5 39 Poor
70-77 4 5 5 10 10 -5 29 Poor
6
8.7-13.7 4 5 5 3 10 -5 22 Poor
23.7-26.7 7 8 8 14 10 -5 42 Fair
26.7-33.7 7 13 10 25 10 -5 60 Fair
8
12-21 2 5 5 1 7 -5 15 Very Poor
21-23 2 5 5 6 10 -5 23 Poor
23-35 2 5 5 10 10 -5 27 Poor
35-61 2 5 5 15 10 -5 32 Poor
Power House
26
264.1-274.4 7 8 15 24 0 0 54 Fair
274.4-294.1 7 8 15 20 0 0 50 Fair
318.1-322.3 7 8 15 26 0 0 56 Fair
322.3-325.7 4 5 5 13 0 0 27 Poor
325.7-329.6 7 8 15 26 0 0 56 Fair
329.6-335.1 7 8 15 23 0 0 53 Fair
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
98
Table 4.6: Calculation of Q index values for Kohala Hydropower Project site
BH #
Depth from
Ground
Surface (m)
RQD
Joint
Set No.
(Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint
Water
Reduction
Factor
(Jw)
Stress
Reduction
Factor
(SRF)
Q Rock
Quality
Dam
8
11.7-20.7 10 12.00 1.00 6.00 1.00 10.00 0.01 Ext. Poor
20.7-22.7 24 14.00 1.00 6.00 1.00 10.00 0.03 Ext. Poor
22.7-34.7 21 20.00 1.00 2.00 1.00 10.00 0.05 Ext. Poor
34.7-60.7 18 20.00 1.00 3.00 1.00 10.00 0.03 Ext. Poor
9
15-26 41 15.00 1.00 6.00 0.66 7.50 0.04 Ext. Poor
26-32 49 12.00 3.00 6.00 0.66 7.50 0.18 Very Poor
33-34 14 15.00 3.00 1.00 0.66 5.00 0.37 Very Poor
34-45.4 41 15.00 3.00 1.00 0.66 5.00 1.08 Poor
45.4-49.4 37 12.00 1.00 2.00 0.66 7.50 0.14 Very Poor
49.4-70 58 15.00 3.00 2.00 0.66 2.50 1.53 Poor
70-71 71 4.00 3.00 1.00 0.66 5.00 7.03 Fair
71-74 72 12.00 3.00 1.00 0.66 2.50 4.75 Fair
74-75 57 4.00 3.00 1.00 0.66 2.50 11.29 Good
75-81 64 12.00 1.00 1.00 0.66 2.50 1.41 Poor
81-84 47 12.00 3.00 2.00 0.66 2.50 1.55 Poor
10
2.9-11.9 19 6.00 1.00 6.00 1.00 10.00 0.05 Ext. Poor
11.9-18.7 36 12.00 1.00 6.00 1.00 5.00 0.10 Ext. Poor
18.7-31 26 15.00 3.00 2.00 1.00 2.50 1.04 Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
99
BH #
Depth from
Ground
Surface (m)
RQD
Joint
Set No.
(Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint
Water
Reduction
Factor
(Jw)
Stress
Reduction
Factor
(SRF)
Q Rock
Quality
31-40.9 12 15.00 3.00 6.00 1.00 10.00 0.04 Ext. Poor
40.9-43.4 10 12.00 2.00 1.00 1.00 2.50 0.67 Very Poor
43.4-51.9 10 12.00 3.00 2.00 1.00 1.00 1.25 Poor
51.9-53.6 18 12.00 3.00 1.00 1.00 1.00 4.50 Fair
53.6-61.9 10 12.00 3.00 2.00 1.00 1.00 1.25 Poor
11
5-6 10 12.00 3.00 1.00 0.66 2.50 0.66 Very Poor
6-7 29 6.00 3.00 1.00 0.66 2.50 3.83 Poor
7-13 31 15.00 3.00 2.00 0.66 2.50 0.82 Very Poor
13-20 15 15.00 1.00 6.00 0.66 10.00 0.01 Ext. Poor
20-21 10 3.00 2.00 1.00 0.66 2.50 1.76 Poor
21-25 10 6.00 1.00 6.00 0.66 10.00 0.02 Ext. Poor
25-30 16 12.00 3.00 2.00 0.66 2.50 0.53 Very Poor
30-34 34 15.00 3.00 1.00 0.66 2.50 1.80 Poor
34-35 41 6.00 3.00 1.00 0.66 2.50 5.41 Fair
35-44 9 12.00 3.00 6.00 0.66 10.00 0.02 Ext. Poor
44-50 35 12.00 3.00 1.00 0.66 5.00 1.16 Poor
50-52 22 9.00 3.00 1.00 0.66 2.50 1.94 Poor
52-56 37 12.00 3.00 2.00 0.66 2.50 1.22 Poor
56-57 13 15.00 1.50 2.00 0.66 5.00 0.09 Ext. Poor
57-63 31 9.00 3.00 2.00 0.66 2.50 1.36 Poor
63-71 33 6.00 3.00 2.00 0.66 5.00 1.09 Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
100
BH #
Depth from
Ground
Surface (m)
RQD
Joint
Set No.
(Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint
Water
Reduction
Factor
(Jw)
Stress
Reduction
Factor
(SRF)
Q Rock
Quality
71-84 29 12.00 3.00 1.00 0.66 5.00 0.96 Very Poor
12
4-8 48 6.00 1.00 6.00 0.66 10.00 0.09 Ext. Poor
8-16 17 12.00 1.00 6.00 0.66 10.00 0.02 Ext. Poor
16-29 24 12.00 1.00 6.00 0.66 10.00 0.02 Ext. Poor
29-36 17 15.00 1.00 6.00 0.66 10.00 0.01 Ext. Poor
36-73 33 15.00 1.00 6.00 0.66 5.00 0.05 Ext. Poor
Desander
1
45-51 37 12.00 3.00 1.00 1.00 2.50 3.70 Poor
54-59 36 12.00 1.00 6.00 1.00 7.50 0.07 Ext. Poor
84-89 20 12.00 3.00 2.00 1.00 1.00 2.50 Poor
104-114.4 30 6.00 2.00 2.00 1.00 1.00 5.00 Fair
114.4-115.8 50 3.00 2.00 2.00 1.00 1.00 16.67 Good
115.8-116.8 34 6.00 3.00 2.00 1.00 1.00 8.50 Fair
2
45-50 10 12.00 1.00 6.00 1.00 10.00 0.01 Ext. Poor
53-58 10 12.00 1.00 6.00 1.00 10.00 0.01 Ext. Poor
85-90 36 12.00 3.00 2.00 1.00 1.00 4.50 Fair
105-110 20 12.00 1.00 6.00 1.00 2.50 0.11 Very Poor
3
48-53 27 12.00 1.00 6.00 1.00 7.50 0.05 Ext. Poor
53-58 23 12.00 1.00 6.00 1.00 7.50 0.04 Ext. Poor
58-61 12 12.00 1.00 6.00 1.00 7.50 0.02 Ext. Poor
85-86 21 12.00 3.00 2.00 1.00 1.00 2.63 Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
101
BH #
Depth from
Ground
Surface (m)
RQD
Joint
Set No.
(Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint
Water
Reduction
Factor
(Jw)
Stress
Reduction
Factor
(SRF)
Q Rock
Quality
86-90 68 9.00 3.00 1.00 1.00 1.00 22.67 Good
104-108.55 46 15.00 3.00 3.00 1.00 1.00 3.07 Poor
108.55-111.6 61 3.00 3.00 3.00 1.00 1.00 20.33 Good
111.6-115 84 9.00 3.00 1.00 1.00 1.00 28.00 Good
15
48-53 43 12.00 3.00 2.00 1.00 2.50 2.15 Poor
56-61 47 12.00 3.00 2.00 1.00 2.50 2.35 Poor
85-89 18 12.00 3.00 2.00 1.00 2.50 0.90 Very Poor
89-90 52 12.00 3.00 2.00 1.00 2.50 2.60 Poor
104-107 49 12.00 3.00 3.00 1.00 1.00 4.08 Poor
107-109 33 12.00 3.00 1.00 1.00 1.00 8.25 Fair
109-110 10 6.00 3.00 1.00 1.00 1.00 5.00 Fair
110-115 40 12.00 3.00 1.00 1.00 1.00 10.00 Fair
Diversion Tunnel
4
22.4-24.4 10 15.00 3.00 1.00 1.00 2.50 0.80 Very Poor
24.4-25.4 10 12.00 2.00 1.00 1.00 2.50 0.67 Very Poor
25.4-27.4 17 6.00 3.00 1.00 1.00 2.50 3.40 Poor
37.4-39.1 28 15.00 3.00 1.00 1.00 2.50 2.24 Poor
39.4-55.4 22 12.00 3.00 1.00 1.00 2.50 2.20 Poor
5
55-56 10 15.00 1.00 6.00 1.00 5.00 0.02 Ext. Poor
56-57 10 15.00 3.00 1.00 1.00 2.50 0.80 Very Poor
57-58 10 15.00 1.00 6.00 1.00 5.00 0.02 Ext. Poor
Continued...
CHAPTER-4 CORRELATIONS BETWEEN VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
102
BH #
Depth from
Ground
Surface (m)
RQD
Joint
Set No.
(Jn)
Joint
Roughness
No. (Jr)
Joint
Alteration
No. (Ja)
Joint
Water
Reduction
Factor
(Jw)
Stress
Reduction
Factor
(SRF)
Q Rock
Quality
58-59.2 10 15.00 1.00 6.00 1.00 7.50 0.01 Ext. Poor
59.2-60.2 10 15.00 3.00 1.00 1.00 2.50 0.80 Very Poor
70-77 10 15.00 1.00 6.00 1.00 1.00 0.11 Very Poor
6
8.7-13.7 27 15.00 1.00 6.00 1.00 10.00 0.03 Ext. Poor
23.7-26.7 19 12.00 2.00 1.00 1.00 2.50 1.27 Poor
26.7-33.7 36 12.00 3.00 1.00 1.00 2.50 3.60 Poor
8
12-21 10 12.00 1.00 8.00 1.00 10.00 0.01 Ext. Poor
21-23 26 13.00 1.00 6.00 1.00 10.00 0.03 Ext. Poor
23-35 14 14.00 1.00 2.00 1.00 10.00 0.05 Ext. Poor
35-61 10 20.00 1.00 2.00 1.00 10.00 0.03 Ext. Poor
Power House
26
264.1-274.4 42 15 3 1 0.66 0.5 11.09 Good
274.4-294.1 71 15.00 3.00 4.00 0.66 0.5 4.69 Fair
318.1-322.3 49 6.00 3.00 1.00 0.66 0.5 32.34 Good
322.3-325.7 40 6.00 1.00 6.00 0.66 0.25 2.93 Fair
325.7-329.6 85 4.00 3.00 2.00 0.66 0.5 42.08 Very Good
329.6-335.1 26 12.00 3.00 1.00 0.66 0.5 8.58 Fair
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
103
Table 4.7: Calculation of RSR values for Kohala Hydropower Project site
BH #
Depth from
Ground
Surface (m)
Ratings for
RSR =
A+B+C
Rock
Quality
General
Area
Geology
(A)
Joint
Pattern,
Direction of
Drive (B)
Groundwater,
Joint
Condition (C)
Dam
8
11.7-20.7 6 7 7 20 Very Poor
20.7-22.7 6 7 9 22 Poor
22.7-34.7 6 7 17 30 Poor
34.7-60.7 6 6 6 18 Very Poor
9
15-26 6 7 6 19 Very Poor
26-32 7 14 11 32 Poor
33-34 12 18 15 45 Fair
34-45.4 12 14 15 41 Fair
45.4-49.4 7 6 9 22 Poor
49.4-70 10 18 12 40 Poor
70-71 12 21 15 48 Fair
71-74 12 19 12 43 Fair
74-75 15 29 22 66 Good
75-81 12 20 12 44 Fair
81-84 15 23 14 52 Fair
10
2.9-11.9 12 10 14 36 Poor
11.9-18.7 15 12 14 41 Fair
18.7-31 12 11 12 35 Poor
31-40.9 8 8 9 25 Poor
40.9-43.4 10 23 15 48 Fair
43.4-51.9 10 20 13 43 Fair
51.9-53.6 10 20 12 42 Fair
53.6-61.9 12 22 15 49 Fair
11
5-6 15 20 12 47 Fair
6-7 10 10 12 32 Poor
7-13 8 7 7 22 Poor
13-20 6 7 6 19 Very Poor
20-21 12 17 12 41 Fair
21-25 6 8 7 21 Poor
25-30 10 17 12 39 Poor
30-34 12 24 12 48 Fair
34-35 12 15 15 42 Fair
Continued...
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
104
BH #
Depth from
Ground
Surface (m)
Ratings for
RSR =
A+B+C
Rock
Quality
General
Area
Geology
(A)
Joint
Pattern,
Direction of
Drive (B)
Groundwater,
Joint
Condition (C)
35-44 10 10 9 29 Poor
44-50 15 19 18 52 Fair
50-52 15 18 15 48 Fair
52-56 12 22 12 46 Fair
56-57 10 13 12 35 Fair
57-63 12 17 15 44 Fair
63-71 15 26 15 56 Fair
71-84 15 22 12 49 Fair
12
4-8 10 8 10 28 Poor
8-16 6 7 8 21 Poor
16-29 7 11 8 26 Poor
29-36 12 17 12 41 Fair
36-73 10 15 10 35 Poor
Desander
1
45-51 15 19 18 52 Fair
54-59 10 16 12 38 Fair
84-89 12 19 15 46 Fair
104-114.4 12 18 12 42 Fair
114.4-115.8 15 25 18 58 Fair
115.8-116.8 12 20 11 43 Fair
2
45-50 10 9 9 28 Poor
53-58 8 9 9 26 Poor
85-90 10 20 12 42 Fair
105-110 12 22 15 49 Fair
3
48-53 10 8 9 27 Poor
53-58 10 9 10 29 Poor
58-61 8 8 7 23 Poor
85-86 12 27 18 57 Fair
86-90 12 26 18 56 Fair
104-108.55 12 25 15 52 Fair
108.55-111.6 12 22 15 49 Fair
111.6-115 12 27 18 57 Fair
15
48-53 10 15 15 40 Fair
56-61 12 21 15 48 Fair
85-89 12 21 15 48 Fair
Continued...
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
105
BH #
Depth from
Ground
Surface (m)
Ratings for
RSR =
A+B+C
Rock
Quality
General
Area
Geology
(A)
Joint
Pattern,
Direction of
Drive (B)
Groundwater,
Joint
Condition (C)
89-90 12 24 18 54 Fair
104-107 15 31 18 64 Good
107-109 12 18 18 48 Fair
109-110 15 24 15 54 Fair
110-115 15 35 18 68 Good
Diversion Tunnel
4
22.4-24.4 10 9 12 31 Poor
24.4-25.4 10 12 12 34 Poor
25.4-27.4 10 19 12 41 Fair
37.4-39.1 12 19 15 46 Fair
39.4-55.4 15 27 18 60 Fair
5
55-56 8 8 9 25 Poor
56-57 12 16 15 43 Fair
57-58 7 7 7 21 Poor
58-59.2 8 9 8 25 Poor
59.2-60.2 12 18 15 45 Fair
70-77 12 13 12 37 Poor
6
8.7-13.7 10 8 12 30 Poor
23.7-26.7 12 21 15 48 Fair
26.7-33.7 12 25 15 52 Fair
8
12-21 10 7 8 25 Poor
21-23 10 10 12 32 Poor
23-35 10 9 12 31 Poor
35-61 10 9 12 31 Poor
Power House
26
264.1-274.4 12 21 15 48 Fair
274.4-294.1 12 25 15 52 Fair
318.1-322.3 12 27 15 54 Fair
322.3-325.7 10 10 12 32 Poor
325.7-329.6 12 29 15 56 Fair
329.6-335.1 12 30 18 60 Fair
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
106
Table 4.8: GSI values for Kohala Hydropower Project site
BH #
Depth from
Ground
Surface (m)
GSI Rock Quality
Dam
8
11.7-20.7 18 Very Poor
20.7-22.7 20 Very Poor
22.7-34.7 25 Poor
34.7-60.7 28 Poor
9
15-26 13 Very Poor
26-32 28 Poor
33-34 39 Fair
34-45.4 45 Fair
45.4-49.4 28 Poor
49.4-70 32 Poor
70-71 49 Fair
71-74 48 Fair
74-75 56 Good
75-81 37 Fair
81-84 46 Fair
10
2.9-11.9 29 Poor
11.9-18.7 31 Poor
18.7-31 29 Poor
31-40.9 27 Poor
40.9-43.4 35 Poor
43.4-51.9 46 Fair
51.9-53.6 51 Fair
53.6-61.9 37 Fair
11
5-6 36 Fair
6-7 42 Fair
7-13 26 Poor
13-20 24 Poor
20-21 46 Fair
21-25 26 Poor
25-30 32 Poor
30-34 30 Poor
34-35 45 Fair
35-44 25 Poor
44-50 33 Poor
50-52 42 Fair
Continued...
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
107
BH #
Depth from
Ground
Surface (m)
GSI Rock Quality
52-56 36 Fair
56-57 30 Poor
57-63 39 Fair
63-71 45 Fair
71-84 37 Fair
12
4-8 26 Poor
8-16 22 Poor
16-29 26 Poor
29-36 28 Poor
36-73 35 Poor
Desander
1
45-51 42 Fair
54-59 29 Poor
84-89 36 Fair
104-114.4 37 Fair
114.4-115.8 46 Fair
115.8-116.8 42 Fair
2
45-50 26 Poor
53-58 24 Poor
85-90 56 Good
105-110 32 Poor
3
48-53 26 Poor
53-58 23 Poor
58-61 18 Very Poor
85-86 32 Poor
86-90 46 Fair
104-108.55 40 Fair
108.55-111.6 42 Fair
111.6-115 41 Fair
15
48-53 39 Fair
56-61 42 Fair
85-89 39 Fair
89-90 40 Fair
104-107 58 Good
107-109 32 Poor
109-110 46 Fair
110-115 46 Fair
Continued...
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
108
BH #
Depth from
Ground
Surface (m)
GSI Rock Quality
Diversion Tunnel
4
22.4-24.4 32 Poor
24.4-25.4 26 Poor
25.4-27.4 39 Fair
37.4-39.1 36 Fair
39.4-55.4 34 Poor
5
55-56 26 Poor
56-57 46 Fair
57-58 28 Poor
58-59.2 22 Poor
59.2-60.2 30 Poor
70-77 34 Poor
6
8.7-13.7 28 Poor
23.7-26.7 39 Fair
26.7-33.7 40 Fair
8
12-21 29 Poor
21-23 40 Fair
23-35 36 Fair
35-61 18 Very Poor
Power House
26
264.1-274.4 19 Very Poor
274.4-294.1 48 Fair
318.1-322.3 46 Fair
322.3-325.7 28 Poor
325.7-329.6 46 Fair
329.6-335.1 46 Fair
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
109
4.5. CORRELATIONS BETWEEN FOUR ROCK CLASSIFICATION SYSTEMS
The data presented in Table 4.1 to 4.8 has been summarized in Table 4.9 giving a range of
the rating values in each of the four (4) systems for both the sites. Standard deviations
have also been reported. Based on the mean values, the classifications of rock mass of the
Diamer Basha Dam site show that the rock is from Good to Very Good in all four
classification systems while for Kohala, the classifications show Poor/Fair quality of rock
as described in the Table 4.9 as following;.
Table 4.9: Summary of rock mass classifications of Basha and Kohala Sites
Site
RMR Q System RSR GSI
Basha
Range of Rating
Values 59 - 87 34 -126 65 - 85 45 - 65
Mean 73 78 75 55
Rock Quality Good Very Good Good Fair/Good
Standard Deviation 7.54 23.51 4.81 5.33
Kohala
Range of Rating
Values 8 - 69 0.01 - 42 13 - 68 13 - 58
Mean 38 3.47 40 35
Rock Quality Poor Poor Poor Poor/Fair
Standard Deviation 16.39 12.58 6.99 9.50
The data presented in the Tables 4.1 to 4.8 has been shown in the form of histograms
representing the frequency of numerical values of different classification systems for
Basha and Kohala sites in Figures 4.4 and 4.5, respectively.
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
110
Figure 4.4: Frequency of four rock classification systems for Diamer Basha Dam site
Figure 4.5: Frequency of four Rock Classification Systems for Kohala site
It can be inferred from Figures 4.4 and 4.5, that generally the data concentration for all
the systems except Q system is from 50 - 90 for Basha and from 20 - 60 for Kohala which
is also the indication of their rock quality. The values in Q system have different limits as
0
5
10
15
20
25
30
35
40
0 10 20 30 40 50 60 70 80 90 100 110 120 130
No
. o
f D
ata
Po
ints
Q
RMR
RSR
GSI
Numerical Value of Q, RMR, RSR and GSI
0
5
10
15
20
25
30
35
40
45
50
0< 10 20 30 40 50 60 70 80
No. of
Dat
a P
oin
ts
Q
RMR
RSR
GSI
Numerical Value of Q, RMR, RSR and GSI
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
111
the system has a wide range for classification i.e. from 0.001 to 1000. Due to this reason,
standard deviation calculated is more in Q system as compared to others. For Kohala,
many Q system values are below zero with minimum calculated value as 0.01. The
combination of data of both sites gives a wide range for study which has a considerable
advantage for regression analyses. Furthermore, the wide rock mass classes also provide
an advantage in the use of the correlations.
The surface mapping done in the Adits of Basha has given more reliable information
about fracture trace length. On the other hand, borehole information gives a continuous
logging of the fracture frequency, fracture surface characteristics and orientation, but less
information about trace length as observed from the borehole data of Kohala site.
It has been observed that RMR and Q systems are the most comprehensive systems to
apply among the four systems used having all the parameters involved related to various
rock properties. Both systems have many factors in common as well as a few factors to
differ. It is also observed that RMR system is easy to apply as compared to Q system. The
lengthy tables make the Q system a bit difficult for users. Therefore, more practice and
familiarization is required to use Q system. More reliable data can generate good
correlations among the systems having different parameters. In this study extreme care
has been taken to cautiously consider all the parameters involved to get the reliable rating
values and a good correlation, consequently.
A total number of 143 (48 of Basha and 95 of Kohala) rating value sets in four
classification systems have been used for analyses as presented in Table 4.1 to 4.8. A
series of regression analyses were performed to obtain empirical relations between the
classification systems applied. In these analyses, linear, exponential, logarithmic,
polynomial and power functions were used separately. The comparison of correlation
coefficients obtained from these equations between various systems is shown in Figure
4.6. The correlation having the highest correlation coefficient between different rock mass
classification systems (shown in dark colour) have been selected.
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
112
Correlation between RMR and Q System
Correlation between RSR and Q System
Correlation between RSR and RMR
Correlation between GSI and RMR
Correlation between GSI and Q System
Figure 4.6: Comparison of correlation coefficients between various systems
0
0.2
0.4
0.6
0.8
1
Lin. Exp. Log. Poly. Power
Co
rrea
ltio
n C
oef
fici
ent
0
0.2
0.4
0.6
0.8
1
Lin. Exp. Log. Poly. Power
Co
rrea
ltio
n C
oef
fici
ent
0
0.2
0.4
0.6
0.8
1
Lin. Exp. Log. Poly. Power
Corr
ealt
ion C
oef
fici
ent
0
0.2
0.4
0.6
0.8
1
Lin. Exp. Log. Poly. Power
Corr
ealt
ion C
oef
fici
ent
0
0.2
0.4
0.6
0.8
1
Lin. Exp. Log. Poly. Power
Corr
ealt
ion C
oef
fici
ent
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
113
Using these numerical values for both the sites, attempts have been made to develop two
types of correlations. First keeping the data separately and second by combining the data
of both the sites. Figure 4.7 shows the different correlations between four classification
systems.
Correlations between Q System and RMR
Correlations between Q System and RSR
Correlations between RMR and RSR
Correlations between RMR and GSI
Figure 4.7: Correlations between various systems using separate data
Two different rocks have yielded slightly different correlations. It is notable that the
correlation coefficients are ranging from 0.495 to 0.821.
RMR = 13.76lnQ + 13.46
R² = 0.592
RMR = 6.274lnQ + 41.49
R² = 0.821
0
10
20
30
40
50
60
70
80
90
100
0.01 1.00 100.00
RM
R
Q
Basha
Kohala
RSR = 7.610lnQ + 41.84
R² = 0.495
RSR = 4.367lnQ + 42.98
R² = 0.676
0
10
20
30
40
50
60
70
80
90
100
0.01 1.00 100.00
RS
R
Q
Basha
Kohala
RSR = 0.514RMR + 37.21
R² = 0.649
RSR = 0.659RMR + 15.48
R² = 0.737
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
RS
R
RMR
Basha
Kohala
GSI = 0.480RMR + 20.46
R² = 0.562
GSI = 0.470RMR + 17.31
R² = 0.6590
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
GS
I
RMR
Basha
Kohala
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
114
By combining the data of both the sites, correlations have been developed between four
(4) classification systems by regression analysis as described in Figures 4.8 to 4.12.
Figure 4.8: Correlation between Q System and RMR
Figure 4.9: Correlation between Q System and RSR
RMR = 6.808lnQ + 42.34
R² = 0.901
0
10
20
30
40
50
60
70
80
90
100
RM
R
Q System
Kohala
Basha
RSR = 5.921lnQ + 45.69
R² = 0.856
0
10
20
30
40
50
60
70
80
90
100
RS
R
Q System
Kohala
Basha
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
115
Figure 4.10: Correlation between RMR and RSR
Figure 4.11: Correlation between RMR and GSI
RSR = 0.839RMR + 10.33
R² = 0.885
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
RS
R
RMR
Kohala
Basha
GSI = 0.535RMR + 15.43
R² = 0.835
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
GS
I
RMR
Kohala
Basha
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
116
Figure 4.12: Correlation between Q system and GSI
In Figures 4.8 to 4.12, lower rating values demonstrate the relatively weak rock mass
(mostly of Kohala site) while higher values correspond to the strong rocks (mostly of
Basha site). The correlation coefficients of the selected functions range between 0.77 and
0.90. This is a quite reasonable range because, while classifying the rock mass systems,
the numerous variations that occur in rock masses and the uncertainties involved in
observing and recording the different parameters can lead to very low regression
coefficients. Moreover, different classification systems place different emphases on the
various parameters. For example stress is not used specifically in RMR, whereas Q
system uses a stress reduction factor. Also material strength is an integral part in RMR
but not in Q system (Milne et al., 1998). So there are always chances of some weak
correlations. All the rock mass classification systems have some limitations, but if applied
appropriately and with care they are valuable tools and can generate very useful
correlations to use. Based on the study, following correlations (Eq. 4.1 to 4.5) have been
proposed;
GSI = 3.690lnQ + 38.05
R² = 0.772
0
10
20
30
40
50
60
70
80
90
100
0.001 0.010 0.100 1.000 10.000 100.000
GS
I
Q System
Kohala
Basha
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
117
𝑅𝑀𝑅 = 6.81 𝑙𝑛𝑄 + 42.34 (4.1)
𝑅𝑆𝑅 = 5.92 𝑙𝑛𝑄 + 45.96 (4.2)
𝑅𝑆𝑅 = 0.84 𝑅𝑀𝑅 + 10.33 (4.3)
𝐺𝑆𝐼 = 0.54 𝑅𝑀𝑅 + 15.43 (4.4)
𝐺𝑆𝐼 = 3.69 𝑙𝑛𝑄 + 38.05 (4.5)
4.6. COMPARISON WITH EXISTING CORRELATIONS
Due to the common usage of classification systems, a number of statistical correlations
have already been developed by various researchers to relate the rock mass rating values
derived from different systems to each others. The correlations suggested in this study can
enable the ground quality to be found directly and independently in any of the four (4)
systems from only one set of observations. Thus, the estimated rock support found in one
system can be easily checked in other systems. This method results in better rock support
estimates, provided that the ground characterization is properly made.
As most of the practiced systems in the rock engineering are RMR and Q system,
therefore most of the literature is found to be on the correlations of these two systems.
The relationship between the RMR and Q is in the form of RMR = A ln Q + B, where A
is generally between 5 and 15 and B is between 36 and 49 (refer Table 2.16, Chapter 2).
Bieniawski (1976) suggested the relationship, RMR = 9 ln Q + 44 which is the most
popular and used equation for correlating the two systems. Rutledge and Preston (1978)
presented correlations between three systems, RMR, Q system and RSR. Similarly Tugrul
(1998) has studied the relations between these three systems and suggested new
correlations.
GSI being relatively new system, therefore less reference is found in the literature about
its relationship with other systems.
The comparisons of correlations developed in this study with the most renowned existing
correlations are presented in graphical form in Figures 4.13 to 4.16.
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
118
Fig. 4.13: Comparison of Correlations between Q System and RMR
Fig. 4.14: Comparison of Correlations between Q System and RSR
0
10
20
30
40
50
60
70
80
90
100
0.001 0.01 0.1 1 10 100
RM
R
Q System
RMR=6.808lnQ+42.34 (This Study)
RMR=9lnQ+44 (Bieniawski, 1976)
RMR=13.5lnQ+43 (Rutledge & Preston, 1978)
0
10
20
30
40
50
60
70
80
90
100
0.001 0.01 0.1 1 10 100
RS
R
Q System
RSR=5.92lnQ+45.69 (This Study)
RSR=13.3lnQ+46.5 (Rutledge & Preston, 1978)
RSR=6lnQ+46 (Tugrul, 1998)
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
119
Fig. 4.15: Comparison of Correlations between RMR and RSR
Fig.4.16: Comparison of Correlations between RMR and GSI
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
RS
R
RMR
RSR=0.839RMR+10.33 (This Study)
RSR=0.77RMR+12.40 (Rutledge & Preston, 1978)
RSR=0.78RMR+17 (Tugrul, 1998)
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
GS
I
RMR
GSI=15.43+0.535RMR (This Study)
GSI=4.71+0.69RMR (Milne et al. 1998)
GSI=RMR-5 (Hoek, 1995)
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
120
The lists of existing correlations between different rock mass classification systems have
been presented in Literature Review (Chapter 2). Each of these expressions has been
derived from a series of specific data taken from some worksites. Therefore use of these
correlations with extreme prudence about the compatibility of the data has been
recommended by many researchers. Bieniawski’s famous equation (Eq. 2.11) is based on
the many case histories having Poor to Good quality rocks. Rutledge & Preston (1978)
has worked on weak sedimentary rocks to develop the correlations (Eq. 2.12, 2.25 &
2.28) while Tugrul (1998) has developed his correlations working on limestone (Eq. 2.23,
2.26 &2.29). The Hoek’s correlation between GSI and RMR (Eq. 2.32) is based on good
competent rocks with GSI > 25 (Zhao, 2010).
While comparing such correlations with each other, it should be kept in mind that mostly
the correlations are based on local geological data. So some variations may always be
expected. The correlations developed in this study are in comparison with the other
existing correlations as the slopes of the graphs demonstrate. There is slight variation in
the correlation of Q system with RSR. The reason may be that Rutledge (1978) has
developed his correlation by using the classifications of weak rocks. Also due to fact that
Q system is relatively difficult to familiarize having large variations in the input
parameters, some deviation can be expected in correlations involving Q system. However
the correlation of Tugrul (1998) is very much similar to the correlation developed in this
study.
It is also observed that while comparing the correlations, the middle portion of graphs in
Figure 4.13, 4.14 and 4.16 are close to each other. This is probably because, for these
portions the rock quality is from fair to good for which jointing parameters are relatively
easy to understand. So less variation may be expected as the application of the systems
will be easy.
4.7. SUMMARY
Rock masses of both the sites have been classified in four (4) major rating systems. The
rock mass rating for the Basha dam site varies in the narrow range in all the systems. For
example it varies from 40 to 120 as determined in Q system. Such a narrow range depicts
that rock mass at Basha dam site is fairly homogenous, whereas for Kohala site, the Q
CHAPTER 4 CORRELATIONS BETWEEN VARIOUS
ROCK MASS CLASSIFICATION SYSTEMS
121
values vary between 0.01 and 40 indicating variable and relatively poor quality rocks as
compared with Basha dam site.
The correlations developed among various rock mass classification systems have good
regression coefficients (from 0.835 to 0.901) indicating good correlations. The
correlations developed through present study are generally in comparison with the other
existing correlations being used across the world. However, some of the existing
correlations do not match with those developed by this study indicating that such
correlations are quite empirical and may only be applied to similar rock type and
conditions. Further, due to fact that Q system is relatively difficult to use having large
variations in the input parameters, some deviation can be expected in correlations
involving Q system. The difficulty when applying borehole data to the RMR or Q system
is that the RQD as one of the main input parameters in both the systems is strongly
direction dependant. In case the borehole was drilled parallel to any major joint or
discontinuity, the RQD would be constantly lower and pretend a much worse condition.
CHAPTER-5
122
CORRELATIONS BETWEEN DEFORMATION MODULUS AND
VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
5.1. INTRODUCTION
The in situ tests to determine the deformation modulus are quite expensive, time
consuming and require special procedures. There have been several attempts to correlate
the modulus with different rock mass classification systems. The first empirical model for
prediction of the modulus of deformation of rock mass was developed by Bieniawski in
1978 which correlates the modulus with RMR. After Bieniawski, some other empirical
approaches were developed with other systems like RSR, GSI and Q system. All
empirical relations are open to the improvement as a result of new data. For their
empirical equation, Hoek (1997) stated that, as more field evidence is gathered, it may be
necessary to modify the relation. This statement is valid for all empirical approaches.
In this research, modulus of deformation determined at Diamer Basha Dam and Kohala
Hydropower Project sites by Plate Load and Flat Jack tests have been correlated with four
main rock mass classification systems i.e. RMR, Q System, RSR and GSI. New equations
have also been proposed for prediction of deformation modulus from the four rock mass
classification systems applied.
5.2. PLATE LOAD TESTS AT DIAMER BASHA DAM SITE
For assessment of the deformation modulus of the rock mass, Plate Load tests were
planned in the Adits of Basha dam. Eight (8) tests in Adit 4 were carried out in 2011 and
2012. Among these, four (4) tests were horizontal and remaining four (4) were vertically
oriented.
Adit 4 was excavated on the left bank of River Indus at the elevation of ±975 m having
total length of 532 m. It has a standard cross section with a width of 2.4 m, a height of 3.2
m and a circular crown. The initial 150 m of main Adit run in southwest direction.
Thereafter the Adit turns northwest (azimuth 300° – 120°). The cross cut starts at
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chainage 316.55 m in northeast direction. The total length of cross cut is 100 m long. A
view of the portal of the Adit 4 is shown in Figure 5.1.
Figure 5.1: View of portal of Adit 4 of Diamer Basha Dam site
From the portal of the Adit the rock mass is massive having complicated joint pattern.
Most of the joints are tight and show no infill. The joint spacing narrows as we go further
in the Adit and at Ch. 0+118 a fault is intersected which is manifested as a closely
fractured zone with a width of 2 m. The rock mass beyond this fault is massive and fresh
but the joint pattern is very diverse and in parts narrowly spaced which leads to closely
fractured portions. The rock mass within the Adit has been classified as good with
portions of fair rock with local extent as described in chapter 4. Out of eight (8) tests
performed in the Adit, four (4) were conducted in the cross cut to observe the possible
effect of anisotropy. Locations for the tests were carefully selected as per ASTM
standards with the two test surfaces nearly parallel and in planes oriented perpendicular to
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the thrust of the loading assembly. Locations of the test sites inside Adit 4 are described
in the Table 5.1.
Table 5.1: Locations of the Plate Load tests in Adit 4 of Basha site
Test
No. Location Orientation Chainage Span (m)
Geological Conditions
1 Main Adit Horizontal 248.5 3.74 25
o/015
o, No major
discontinuity
2 Main Adit Vertical 238.4 3.38 Dip > 45
o, Joint striking
subparallel to cavern axis
3 Main Adit Horizontal 391.5 3.99 Dip < 45
o, Joint striking
subparallel to cavern axis
4 Main Adit Vertical 393.0 4.0 26o/028
o, Pegmatite dyke
5 Right X-Cut Horizontal 16.0 3.57 Dip > 45
o, Joint striking
subparallel to cavern axis
6 Right X-Cut Vertical 18.0 3.85 Dip > 45
o, Joint striking
subparallel to cavern axis
7 Right X-Cut Vertical 91.0 4.0 No major discontinuity
8 Right X-Cut Horizontal 94.0 3.5 Dip < 45
o, Joint striking
subparallel to cavern axis
5.2.1. Equipment Used
The equipment used for the tests have been divided into four (4) groups as follows;
A. Equipment related to installation of rigid plates.
B. Hydraulic system for applying hydraulic pressure
C. Scaffolding for erecting the spacers, plates and Flat Jacks etc.
D. Equipment relating to measurements through extensometers and data loggers
All the equipment were carefully checked and calibrated before the execution of the tests.
The size of the plate is usually determined by local geology, pressures to be applied, and
the size of the Adit to be tested. Recommended plate diameter is commonly 0.5 to 1 m.
For Basha, plates of 0.9 m dia were used. Larger plates and higher loads measure the
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response of rock further away from the test Adit and hence in situ undisturbed modulus
can be determined. The detail of equipment used is given in the Table 5.2.
Table 5.2: Detail of equipment and accessories used in the Plate Load tests at Basha
Sr.
No.
Description Quantity Remarks
DETAIL OF EQUIPMENT FOR GROUP “A”
1. Rigid Plates of 900 mm dia,
Thickness 25mm
04 Two (02) on each side within the Adit
2. Rigid Plates of 700 mm dia,
Thickness 25mm
02 One on each side. Towards the spacers
side
3. Spacers having length of
1000 mm
08 All spacers are of equal diameter
(approximately 15 mm) provided with
the male and female parts to inter lock
4. Spacers having length of 750
mm, 500 mm, 250 mm, 100
mm & 50 mm
04 each
5. Bracing Plates 5 mm thick 08 Bracing plates are provided with the 04
holes at equal spaces. The bracing
plates bind the spaces and stabilize the
assembly.
6. Adjustable spacers 04 These spacers are used to eliminate all
loose spaces.
7. Key 01 For tightening the system to eliminate
all spaces between spacers, plates and
Jacks
8. Level 01 For levelling the mortar and the
assembly
9. Cement mortar To apply on the rock surface, so that
plates may be installed in a levelled
position.
10. Tool box Large tool box with all the tools
required for this type of test.
DETAIL OF EQUIPMENT FOR GROUP “B”
1. Hydraulic Pump 02 Capacity 250 bars (25 MPa) with slow
unloading facility.
2. Pump leads 05 Sufficient capacity to bear the required
pressure of 9 MPa.
3. Jack with 9 inch Ram length 01 To lift the rigid plates and Flat Jacks to
a height up to the roof in the final stage
of installing the plates.
4. Hydraulic Oil To fill the Flat Jacks to generate
pressure.
5. Flat Jacks of 900 mm dia 02 Sufficient capacity to bear up to 9 MPa
Continued...
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Sr.
No.
Description Quantity Remarks
pressure with a hole of 76 mm at the
centre to pass the extensometer cables.
DETAIL OF EQUIPMENT FOR GROUP “C”
1. Steel pipes of 50 mm dia
having length of 03 m, 02 m
and 01 m.
Scaffolding for installation of spacers,
bracing plates, rigid plates, Flat Jacks,
during both horizontal and vertical tests
2. Keys 04 For erecting the Scaffolding
3. Wrench 04 For lifting the rigid plates and jacks to
a required height.
DETAIL OF EQUIPMENT FOR GROUP “D”
1. Retrievable borehole
extensometers (BOF-EX)
02
(10 rods)
50 mm Stem, with cables sufficiently
large to come out of the boreholes and
to connect with the data logger
2. End stoppers 02 To place at the end in the borehole
where first extensometer rod is to be
installed
3. Extension Rods Different lengths to install the
extensometers at different depths
4. Centralizers 04 To keep the extensometers including
extension rods in the aligned position
5. Stoppers 12 To install the extensometers at different
locations in a borehole. The stoppers
are provided to pass the cables of
already installed extensometers out of
boreholes to connect with the data
logger.
6. Spanners with extension
pipes
05 For screwing the stoppers of the
extensometers.
7. Data loggers with dry battery
and charger
02 Data logger model CR 850 Roctest of
Canada, .having capability of storing
the data at required time intervals.
5.2.2. Methodology
Tests were performed as per ASTM Designation D4394 and D4395. The method is based
upon the measurement of the deformations inside the rock surface which is subjected to
loading as shown in Figure 5.2.
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Figure 5.2: Set up of Plate Load test at Diamer Basha site
The test is normally performed at ambient temperature, but equipment can be modified or
substituted for operations at other temperatures. Some alterations were made at the site
e.g. tunnel diameter gauge and particle board pads were not used in the tests. Following
steps were performed for testing;
Drilling of Boreholes and Surface Preparations
i. Boreholes of 76 mm dia and 6 meter depth were drilled in the opposite parallel
faces of the rock surface (horizontal or vertical) on all the selected locations. The
test locations were carefully chosen within the Adit and relatively less disturbed
zones were selected.
ii. The cores obtained from the boreholes were examined and logged.
iii. Rock surfaces of dimension 1.5 m x 1.5 m on opposite faces were made smooth
with the help of diamond cutter and chiselling. The surfaces were washed with
water to remove any loose particles. The surfaces were prepared in such a way
that the boreholes already drilled were at the centre of the area.
iv. A layer of cement mortars was applied on the surfaces not more than 1.5 inches
thick to make the surface totally smooth and making the opposite faces completely
parallel.
v. Joined the Flat Jacks of 900 mm dia in series with each other through hydraulic
pipes and filled the Jacks with the oil by hydraulic pumps.
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Extensometers Installation
i. The extensometers were installed in the boreholes by following the ASTM
Designation D4403.
ii. First of all, end stopper was installed at approximately 6 meter depth in both the
opposite boreholes with the help of pipes and spanners.
iii. First extensometer was assembled and inserted in the borehole with the help of
extension pipes and centralizer by connecting the wires with the data logger.
iv. The cable was marked with the identification mark as No. 1 –T (Top) or B
(Bottom), or L (Left), or R (Right) according to the position.
v. The second extensometer was assembled with extension pipes and stoppers. The
cable of first extensometer was passed through the hole provided with the stopper
of second extensometer. The cable of second extensometer was attached with data
logger in the same manner as before. The identification mark was placed on the
second cable.
vi. Similarly all the five extensometers were installed one by one, by passing the
cables of already installed extensometers from the lock system of the
extensometer which was going to be installed. During the process, the cables of
the extensometer were kept connected with the data logger to watch the reading of
the extensometer and the position of the sensor on the data logger.
vii. Same process was adopted to install the extensometers on the opposite side of the
rock surface.
viii. Two sets of cables (five from both opposite sides) were tied with paper tape
separately to avoid confusion while connecting with the data logger for recording.
Installation of Jacking System
i. The first rigid plate of 900 mm dia was placed on the wall such that the hole of the
plate exactly coincided with borehole and cables were pulled out through the rigid
plate. The Flat Jack was then placed against the rigid plate such that the hole at the
centre of the jack coincided with the hole of installed rigid plate and also the
cables pulled through the hole of the Flat Jack. After that, another rigid plate of
900 mm was placed on the Flat Jack with extra reinforcement of 700 mm rigid
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plate. So the extensometer cables got out through the plate of 900 mm dia, Flat
Jack of 900 mm dia, then two plates of 900 mm and 700 mm dia respectively.
ii. Same set was also placed on the opposite face of wall for horizontal test and top or
bottom for vertical test. Spacers were installed in such a way that the nozzle of
first spacer enters in the slot of the next spacer through bracing plates. The bracing
plates were used to prevent the system from slipping and gave extra support for
four lines of spacers and stabilized the system.
iii. When the four lines of spacers reached on the other side against the plates and
jack system, adjustable spacers were placed to tight up the system.
iv. While tightening the system, nozzles of the Flat Jacks were kept open so that extra
oil from both the jacks may come back in to the pump.
v. The cables 1, 2, 3, 4 & 5 from both sides were connected with the data logger.
Adjustment of Data Logger
i. The time in the data logger was adjusted by fixing the scanning time and
triggering time in the logger as 10 minutes interval. Normally scanning time and
triggering time was kept same as recommended in the manual of logger.
ii. The test ID was entered in the logger.
iii. Applied pressured was entered according to the loading and unloading cycles. The
peak pressure was taken as 9 MPa which is more than twice the overburden
pressure at the test site; the average overburden at the site is 116 m thick with
average density of the rock as 2.9 g/cc. The peak pressure was divided into 5
cycles and each cycle was further divided into further 10 increments in which 5
were for loading and 5 were for unloading as described in the Table 5.3.
iv. To start with the test, the load was entered as zero (0) MPa and waited for 10
minutes to record the zero reading in the logger. The readings of all the sensors in
the data logger could be viewed through view scan.
v. First load of cycle 1 (0.4 MPa) was applied by increasing the load through pump
in the Flat Jack. The data logger was started and waited for twenty minutes so that
two readings may be stored in the logger (logger had been fixed for 10 minutes
interval).
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Table 5.3: Sequence of applied pressure in Plate Load tests at Basha site
Cycle No. Loading /
Unloading Incremental Pressure (MPa)
1 Loading 0.4 0.8 1.2 1.6 2.0
Unloading 1.6 1.2 0.8 0.4 0
2 Loading 0.8 1.6 2..4 3.2 4.0
Unloading 3.2 2.4 1.6 0.8 0
3 Loading 1.2 2.4 3.6 4.8 6.0
Unloading 4.8 3.6 2.4 1.2 0
4 Loading 1.5 3.0 4.5 6.0 7.5
Unloading 6.0 4.5 3.0 1.5 0
5 Loading 1.8 3.6 5.4 7.2 9.0
Unloading 7.2 5.4 3.6 1.8 0
vi. After recording two readings, the load was increased to second step of cycle 1 (0.8
MPa). Again waited for twenty minutes to record two more readings. Similarly the
load was kept on increasing as per the Table 5.3 and waited for twenty minutes for
each incremental load to record two readings.
vii. The load was decreased step by step as stated in the Table 5.3, till zero (0) by
staying 20 minutes on each step. In this way first step was completed.
viii. The 2nd cycle of 4 MPa pressure was started and proceeded in the same way as in
the first cycle but with the loading and unloading steps as described in Table 5.3.
ix. Similarly all the five cycles were completed by loading and unloading each cycle.
x. After completing the test, the data logger was disconnected. The data was
transferred to PC through data logger software DL-1600, which is in Window
based Excel format.
xi. The data was scrutinized and plotted by choosing one reading for each step and
modulus of deformation was calculated as explained in ASTM Designation
D4395.
Photographs taken during the execution of the tests are shown in Figure 5.3 to 5.8.
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Figure 5.3: Drilling in the floor of the Adit to install multipoint borehole extensometer
Figure 5.4: Rock surface preparation and installation of extensometer
Drilling for Multipoint
Borehole Extensometer
Prepared area (1.5 m x 1.5 m)
Hole for MPBX
Cable of MPBX
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Figure 5.5: Different accessories used in Plate Load test
Spacers of different
lengths
Plates of different
diameters
Bracing plates
and flat jacks Spacers
Cable Stopper
Centerlizer Data Logger
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Figure 5.6: Flat jack and hydraulic pump
Figure 5.7: Installation of plates, Flat Jacks and spacers
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Figure 5.8: Final set up of equipment before load application
5.2.3. Determination of Modulus of Deformation
As described in Table 5.3, the load was applied in five (5) successive loading and
unloading cycles. The deformations were recorded at each sensor of Multipoint Borehole
Extensometer. The arrangement to determine the deformation at the surface could not be
made; therefore in each test ten (10) load-deformation relations were obtained.
The load-deformation curves of all the tests have been drawn and included in Appendix-
E. Depending upon the orientation of the discontinuities with respect to loading, the
shapes of the curves have been formed. In some cases it is a typical smooth loading-
unloading curve. However, in many cases the shape of the curve is slightly irregular
which may be due to open joints and microfractures. The results show that the permanent
deformations are mainly caused by the constant stress that is applied for a period of time
at the peak of the loading–unloading cycles. These deformations can be attributed to
creep-like behaviour of the rock mass.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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The modulus of deformation was calculated from the deflection at a point within the rock
mass beneath the centre of an annularly loaded area described in ASTM Designation
D4395, as follows:
𝐸 𝑚 =2𝑄 1 − 𝜈2
𝑊𝑧 𝑅2
2 + 𝑍2 1 2 − 𝑅12 + 𝑍2 1 2
+𝑍2𝑄 1+𝜈
𝑊𝑧 𝑅1
2 + 𝑍2 −1 2 − 𝑅22 + 𝑍2 −1 2 (5.1)
Where;
𝜈 = Poisson’s ratio of the rock, (average value from laboratory tests, used as 0.25)
𝑄 = Peak pressure on loaded area, (9 MPa),
𝑍 = Depth beneath center of loaded area,
(varying from 0.5 to 6.0 m),
𝑊𝑧 = Deflection at depth Z, (recorded
from sensors in mm at each load interval).
𝑅2 = Outside radius of bearing plate, (450
mm), and
𝑅1 = Inside radius of bearing plate, (38 mm).
The equation (5.1) is based on the elastic solution for uniformly distributed load over
circular area acting on a semi infinite isotropic medium. The deflection is defined as the
movement in the direction of applied load. The equation does not include the stress
history of the rock.
For ease in calculation, Eq. 5.1 has been divided into two parts as follows;
𝐸𝑚 = 𝐾1 +𝐾2 (5.2)
Where;
𝐾1 =2𝑄 1−𝜈2
𝑊𝑧 𝑅2
2 + 𝑍2 1 2 − 𝑅12 + 𝑍2 1 2 (5.3)
and 𝐾2 =𝑍2𝑄 1+𝜈
𝑊𝑧 𝑅1
2 + 𝑍2 −1 2 − 𝑅22 + 𝑍2 −1 2 (5.4)
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Time-rate of loading has negligible influence on the modulus; however the rate was kept
constant during the test. The calculation for modulus is summarized in the Table 5.4.
Table 5.4: Calculation for the Modulus of Deformation at Basha site
Z (mm) Wz (mm) K1 K2 Em (MPa) Em (GPa)
Test No.1 (Horizontal)
Left Side
500 0.122585 23572.65 11647.35 35220.00 35.22
1000 0.057228 28267.51 17172.77 45440.28 45.44
2000 0.019630 42672.37 27749.38 70421.75 70.42
4000 0.009820 43050.87 28519.38 71570.25 71.57
6000 0.005080 55577.89 36947.42 92525.31 92.53
Right Side
500 0.435540 6634.67 3278.22 9912.89 9.91
1000 0.173200 9340.07 5674.18 15014.25 15.01
2000 0.029750 28156.59 18309.93 46466.52 46.47
4000 0.010820 39072.05 25883.58 64955.62 64.96
6000 0.008880 31794.56 21136.58 52931.14 52.93
Test No.2 (Vertical)
Top Side
500 0.268870 10747.44 5310.36 16057.79 16.06
1000 0.079400 20374.09 12377.45 32751.54 32.75
2000 0.034271 24442.29 15894.56 40336.85 40.34
4000 0.015500 27274.81 18068.41 45343.21 45.34
6000 0.010506 26874.16 17865.57 44739.73 44.74
Bottom Side
500 0.146206 19764.35 9765.65 29530.00 29.53
1000 0.044127 36660.03 22271.30 58931.33 58.93
2000 0.021160 39587.00 25743.00 65330.00 65.33
4000 0.008918 47405.72 31404.28 78810.00 78.81
6000 0.005910 47772.41 31758.44 79530.85 79.53
Test No.3 (Horizontal)
Left Side
500 0.136413 21183.26 10466.74 31650.00 31.65
1000 0.064190 25201.79 15310.32 40512.11 40.51
2000 0.026162 32018.63 20821.37 52840.00 52.84
4000 0.011191 37775.40 25024.60 62800.00 62.80
Continued...
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137
Z (mm) Wz (mm) K1 K2 Em (MPa) Em (GPa)
6000 0.007160 39431.83 26213.74 65645.57 65.65
Right Side
500 0.182248 15855.65 7834.35 23690.00 23.69
1000 0.067862 23838.12 14481.88 38320.00 38.32
2000 0.030275 27667.88 17992.12 45660.00 45.66
4000 0.011236 37625.02 24924.98 62550.00 62.55
6000 0.007265 38863.85 25836.15 64700.00 64.70
Test No.4 (Vertical)
Top Side
500 0.193002 14972.18 7397.82 22370.00 22.37
1000 0.078612 20578.42 12501.58 33080.00 33.08
2000 0.028109 29800.84 19379.16 49180.00 49.18
4000 0.012270 34455.01 22824.99 57280.00 57.28
6000 0.007545 37422.22 24877.78 62300.00 62.30
Bottom Side
500 0.170040 16994.02 8396.82 25390.84 25.39
1000 0.080785 20024.77 12165.23 32190.00 32.19
2000 0.035112 23856.43 15513.57 39370.00 39.37
4000 0.013008 32500.08 21529.92 54030.00 54.03
6000 0.005266 53616.49 35643.51 89260.00 89.26
Test No.5 (Horizontal)
Left Side
500 0.186338 15507.62 7662.38 23170.00 23.17
1000 0.070416 22973.43 13956.57 36930.00 36.93
2000 0.024393 34340.07 22330.98 56671.05 56.67
4000 0.010205 41426.62 27443.38 68870.00 68.87
6000 0.006091 46353.87 30815.41 77169.28 77.17
Right Side
500 0.147052 19650.57 9709.43 29360.00 29.36
1000 0.058754 27533.28 16726.72 44260.00 44.26
2000 0.023586 35514.99 23095.01 58610.00 58.61
4000 0.011584 36494.16 24175.84 60670.00 60.67
6000 0.007402 38143.04 25356.96 63500.00 63.50
Test No.6 (Vertical)
Top Side
500 0.278187 10387.49 5132.51 15520.00 15.52
1000 0.075006 21567.53 13102.47 34670.00 34.67
2000 0.031957 26212.39 17045.64 43258.03 43.26
4000 0.012808 33007.10 21865.81 54872.91 54.87
Continued...
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138
Z (mm) Wz (mm) K1 K2 Em (MPa) Em (GPa)
6000 0.008537 33073.32 21986.68 55060.00 55.06
Bottom Side
500 0.168453 17154.09 8475.91 25630.00 25.63
1000 0.086567 18687.30 11352.70 30040.00 30.04
2000 0.030144 27788.42 18070.51 45858.93 45.86
4000 0.009062 46653.82 30906.18 77560.00 77.56
6000 0.006501 43429.00 28871.00 72300.00 72.30
Test No.7 (Vertical)
Top Side
500 0.214799 13452.87 6647.13 20100.00 20.10
1000 0.068488 23620.39 14349.61 37970.00 37.97
2000 0.031415 26663.91 17339.25 44003.16 44.00
4000 0.012124 34869.90 23099.83 57969.74 57.97
6000 0.008089 34901.80 23202.23 58104.03 58.10
Bottom Side
500 0.203037 14232.22 7032.20 21264.42 21.26
1000 0.112721 14351.40 8718.60 23070.00 23.07
2000 0.036426 22995.97 14954.03 37950.00 37.95
4000 0.010763 39279.20 26020.80 65300.00 65.30
6000 0.006805 41491.54 27583.01 69074.55 69.07
Test No.8 (Horizontal)
Left Side
500 0.256077 11284.35 5575.65 16860.00 16.86
1000 0.068075 23763.47 14436.53 38200.00 38.20
2000 0.031614 26496.50 17230.38 43726.88 43.73
4000 0.010306 41021.63 27175.10 68196.73 68.20
6000 0.007203 39199.49 26059.28 65258.78 65.26
Right Side
500 0.158091 18278.51 9031.49 27310.00 27.31
1000 0.074193 21803.92 13246.08 35050.00 35.05
2000 0.036994 22642.94 14724.46 37367.40 37.37
4000 0.010804 39130.84 25922.52 65053.36 65.05
6000 0.006080 46439.21 30872.15 77311.36 77.31
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5.2.4. Variation in Modulus of Deformation
A graph representing the deformations and modulus with respect to corresponding depth
below rock surface in all the tests is shown in Figure 5.9 below.
Figure 5.9: Uniaxial deformations and modulus vs distance in all the tests at Basha site
It is observed that the scatter of data is more near to rock surface and as the depth
increases, the deformations become minute and less scattered. Below 2 m, the
deformations are almost negligible upon applied pressure. Consequently the modulus of
deformation is increasing with the increase in distance from the rock surface.
Likewise, variations in modulus of deformation with respect to distance from rock surface
in each test have been plotted in Figure 5.10.
It is noted conclusively that the modulus is not constant with depth and rather increases
with increase in depth in all the tests. The reason is the low deformation level at increased
depth and secondly the rock has less micro-openings. As the compactness of the rock
increases with depth, deformations recorded decrease and modulus increases. The effect
of blast damage also varies from place to place.
0
10
20
30
40
50
60
70
80
90
100
0.00
0.05
0.10
0.15
0.20
0.25
0.30
0.35
0.40
0.45
0.50
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Def
orm
ati
on
(m
m)
Distance from Rock Surface (m)
Deformation vs Distance Modulus vs Dastance
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Figure 5.10: Variation in the modulus of deformation in each of the tests at Basha site
0
20
40
60
80
100
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No. 1 (Horizontal)
Left
Right0
20
40
60
80
100
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No. 2 (Vertical)
Top
Bottom
0
10
20
30
40
50
60
70
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No. 3 (Horizontal)
Left
Right
0
20
40
60
80
100
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No. 4 (Vertical)
Top
Bottom
0
20
40
60
80
100
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No.5 (Horizontal)
Left
Right
0
20
40
60
80
100
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No.6 (Vertical)
Top
Bottom
0
20
40
60
80
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No.7 (Vertical)
Top
Bottom
0
20
40
60
80
100
0 1 2 3 4 5 6 7
Em
(G
Pa
)
Distance from Surface (m)
Test No.8 (Horizontal)
Left
Right
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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141
However, somewhere slightly erratic trend is also observed which is due to the presence
of joints or a weak zone in the rock mass e.g. in test 1 (right), 6 (bottom) and 8 (left). The
presence of such zones is always expected due to the variance in local deformational
characteristics. Also the installation of extensometer sensors requires extreme care,
negligence in which may result in diverse readings.
The combined graph including all the points as described in Figure 5.10 is shown in
Figure 5.11
Figure 5.11: Variation in modulus of deformation in all the tests at Basha site (combined)
The scatter of data is clear in Figure 5.12 which is always expected in rock testing. The
modulus depends upon the compactness or quality of the rock in terms of classification
ratings (RMR, Q system values etc.). As explained earlier, the modulus of deformation is
generally increasing with increase in distance from the rock surface in all the tests. This
increase in modulus is continuous until a distance is reached after which the modulus is
relatively constant. This distance is about 2 to 3 m from the rock surface i.e. about 2 to 3
times the dia of the plate (0.9 m). Beyond this distance, the effect of loading pressure is
negligible and hence the variation in modulus is also very small.
0
10
20
30
40
50
60
70
80
90
100
0 1 2 3 4 5 6
Em
(GP
a)
Distance from Surface (m)
1-L
1-R
2-T
2-B
3-L
3-R
4-T
4-B
5-L
5-R
6-T
6-B
7-T
7-B
8-L
8-R
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Strains have also been computed from the uniaxial deformations up to 3 m from the rock
surface for the plate dia of 0.9 m and plotted against modulus in Figure 5.12.
Figure 5.12: Variation in modulus of deformation with respect to strain at Basha site
For lower strain values the modulus is higher and as the strain level increases (near the
the rock surface) the modulus values decrease.
5.2.5. Average Modulus of Deformation
As discussed, the influence of measuring location on the modulus is evident from the
data. Therefore the use of an average value of modulus is very important for the designer.
The use of an average value of modulus is advisable because for values very high, the
designer has to apply some safety factor. On the other hand the use of minimum value of
the modulus could result into over conservative design.
The statistics including the range, mean modulus of deformation and standard deviation is
shown in Table 5.5.
0
10
20
30
40
50
60
70
80
0.0000 0.0020 0.0040 0.0060 0.0080 0.0100 0.0120
E (
GP
a)
Strain (%)
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143
Table 5.5: Summary of Modulus of Deformation at Basha site
Em (GPa)
at 0.5 m at 1.0 m at 2.0 m at 4.0 m at 6.0 m
Range 9.91 – 35.22 15.01 – 58 93 37.37 – 70.42 45.34 – 78.81 44.74 – 92.53
Mean 23.31 36.13 49.16 63.33 68.15
Standard
Deviation 6.64 9.65 9.75 8.67 12.88
Considering all measuring points in all the tests, following values are determined;
Mean Em (in all directions) = 47.90 GPa
Mean Em (in horizontal direction) = 49.66 GPa
Mean Em (in vertical direction) = 46.13 GPa
The results are reasonable for the very good quality rock of Gabbronorite at Basha site, as
found in the literature.
It is noted that the modulus is somewhat less in vertical direction as compared to
horizontal direction. The possible reason of this fact is the gravity. The cracks and
discontinuities tend to open due to gravity in vertical direction and hence reduce the
modulus. It is also observed that the joint orientation at some of the test location is near
horizontal and the joints are clay filled. These joints are persistent and are more than
inclined or vertical. Therefore, it was expected that modulus in vertical direction will be
slightly lesser due to more pronounced relaxation of the rock in vertical direction.
The average values of deformation modulus at each measuring points have been
computed and placed at a single location to observe the variation in deformation modulus
with depth. Contours of modulus have been drawn by using “Surfer” software (Figure
5.13). The gradual increase in modulus for first 2 to 3 m is evident. After which the
contours are widely spaced showing slow/gradual change in modulus till the last
measuring point at 6 m from the rock surface of the Adit.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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144
Figure 5.13: Contours of Modulus of Deformation at Basha site
5.3. PLATE LOAD & FLAT JACK TESTS AT KOHALA HYDROPOWER
PROJECT SITE
5.3.1. Geology of Adit 2
Four (4) Plate Load tests were carried out in the Adit 2 of Kohala Hydropower Project
site. The Adit was excavated to physically observe the geology of the power house area
and to conduct the in situ rock mechanics tests. Three types of rocks are encountered in
- Mean value of Em in GPa
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145
Adit 2 i.e. SS-1, SS-2 and Shale. Mud stone is also observed somewhere, but it is
considered as Shale due to its high clayey content. SS-1 is dominant rock unit
encountered during excavation. It is mostly well compacted and hard. The exposed
surface area of SS-1 is about 51.02% of total exposure. At places SS-2 is also observed in
the Adit. The rock unit is very weak. The exposed surface area of SS-2 is about 13.41%
of total exposure. Shale is the second major rock unit exposed in the Adit. The exposed
surface area of shale is 35.56% of total exposure.
The rocks in the Adit are highly jointed. Most joints are random. Only a few joint sets are
measurable. Multiple shear zones are observed in the Adit indicating the presence of
stresses in the area. The shear zones are mostly filled with clay. A continuous joint set
pattern is observed, which shows that the area is sandwiched between two major thrust
i.e. HFT (Himalayan Frontal Thrust) and MBT (Main Boundary Thrust).
5.3.2. Plate Load Tests
Out of the four tests carried out, two (2) tests were vertical while two (2) tests were
carried out in horizontal direction. ASTM Designation D4394 was followed for the
methodology and a rigid plate of 1 ft. (0.305 m) dia was used for the tests. Although the
exact design and materials of rigid plate may differ, the stiffness should at least be the
minimum stiffness necessary to produce no measurable deflection of the plate under
maximum load as per ASTM. Peak load was selected as 50 Tons (490.3 KN) according to
the overburden at the site. Three (3) dial gauges on each plate having equal
circumferential distance were used to record the deformation during the tests. The sites
were carefully selected on the basis of less disturbed areas and the two opposite faces in
each test were smoothened by grinding and chiseling. A mortar pad and rigid metal plate
were installed against each face and the hydraulic loading system was placed between the
rigid plates for cyclic loading and unloading. The modulus is determined using an elastic
solution for a uniformly distributed load (uniform stress) over a circular area acting on a
semi-infinite elastic medium.
Figure 5.14 shows the sketch of the test set up including major equipment.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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146
Figure 5.14: Set up of Plate Load test at Kohala HPP site
Further details of the tests are as shown in Table 5.6.
Table 5.6: Details of Plate Load tests at Kohala site
Test No. Location Orientation Material
1 RD 50 Horizontal SS-I
2 RD 138 Vertical SS-II, Dry Shale
3 RD 145 Horizontal SS-I
4 RD 170 Vertical SS-I
The modulus of deformation was calculated according to the following equation;
E = 1 − ν2 P
2RWa (5.5)
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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147
Where
ν = Poisson’s ratio of the rock (average value from laboratory tests, used as 0.20)
P = Maximum load on the rigid plate, (490.3 KN),
Wa = Average deflection of the rigid plate, (recorded from three deflection gauges in mm),
and
R = Radius of the rigid plate, (150 mm).
Figure 5.15 shows the execution of Plate Load test in the Adit of Kohala.
Figure 5.15: Plate Load test in Adit 2 of Kohala Hydropower Project
The maximum load of 50 Tons was divided into 5 cycles (10, 20, 30, 40 & 50 Tons).
Each cycle was further divided into 5 incremental steps of loading and 5 steps of
unloading. The rate of loading and unloading was kept about 1 minute per increment. The
deflections were recorded at each incremental loading and mean deformations on each of
the two rock faces were calculated.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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A typical load vs deformation curve is presented in Figure 5.16.
Figure 5.17: Typical load vs deformation curve for Plate Load test at Kohala site
Resultantly, by using the equation 5.5, the moduli of deformation calculated in these tests
are shown in Table 5.7.
Table 5.7: Plate Load test results at Kohala site
Test
No. Orientation
Average Deflection
Wa (mm) Em (GPa)
Left / Top Right / Bottom Left / Top Right / Bottom
1 Horizontal 0.23 0.14 6.80 11.02
2 Vertical 0.11 0.36 14.57 4.35
3 Horizontal 0.22 0.25 7.29 6.32
4 Vertical 0.30 0.36 5.27 4.32
0
10
20
30
40
50
60
0 0.5 1 1.5 2 2.5
LO
AD
( T
ON
S)
DISPLACEMENT ( mm)
Plate Load Test Adit No 2 RD-50 ( Horizontal)
Loading Cycles
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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149
The moduli recorded in these tests are quite lower than the modulus at Basha site. The
reason is that weak sandstone is present at Kohala site with lot of discontinuities. Also the
measurements were made only on the surface which may be affected by the blasting of
the Adit.
5.3.3. Flat Jack Tests
Two (2) Flat Jack tests were carried out at Kohala site in the same Adit (Adit 2) to
determine the in situ modulus of deformation. ASTM Designation D4729-04 was
followed for the equipment and methodology.
The in situ stress in the rock mass was relieved by cutting a slot into the rock
perpendicular to the surface of the Adit. The deformation caused by this stress relief was
measured. A hydraulic Flat Jack was placed into the slot and the slot was grouted. A high-
early strength, non-shrink mortar was used for grouting. The mortar included up to 50 %
clean sand by weight, with grain size between 20- and 60-mesh. Clean, potable water was
used for the mortar. As per ASTM, the cured mortar should have strength greater than the
stress applied by the Flat Jack.
The pressure was applied to such extent that the above-measured displacement was
canceled. The modulus of deformation (Em) of the rock mass was evaluated by
incrementally loading the Flat Jack and measuring the deformations at selected points by
using the equation 5.6 below;
E = PLR2π∆Y (5.6)
Where:
P = Pressure in Flat Jack (MPa)
L = Distance between measuring points (500 mm)
R = Stress Distribution Factor
∆Y = Deformation between measuring points (mm)
The stress distribution factor (𝑅) is calculated by the following equation
R = Aq + sin Aq − υ Aq + sin Aq + Az + sin Az − υ Az − sin Az (5.7)
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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150
Where
𝜐 = Poisson’s ratio of the rock mass (average value taken as 0.2)
Aq , Az = Angles, in radians, between the measuring points and the edges of the Flat Jack,
as shown in Figure 5.17 For the tests at Kohala following values were taken.
Aq= 1.571 rad
Az= 0.925 rad
Figure 5.17: Geometric terms in Flat Jack test (ASTM Designation D4729-04)
The results obtained in both the tests are shown in Table 5.8.
Table 5.8: Calculation for the modulus from Flat Jack tests at Kohala site
Test
No.
Location and
Material
Pressure P
(MPa) L (mm) R ∆Y (mm)
Em
(GPa)
1 Chamber 2
SS-I, Mudstone 8.85 500 0.358 0.040 6.30
2 RD 177
SS-II, Shale 8.23 500 0.358 0.051 4.60
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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Execution of Flat Jack test in the Adit is shown in Figure 5.18.
Figure 5.18: Execution of Flat Jack test in the Adit 2 of Kohala HPP.
It is noted that the modulus obtained in the Flat Jack tests are comparable with the
modulus obtained in the Plate Load tests performed at Kohala site. The sandstone is not
very strong; therefore low values of deformation modulus have been obtained in both the
tests.
5.4. CORRELATIONS OF MODULUS OF DEFORMATION
At Basha site, cores extracted from the holes drilled for Multipoint Borehole
Extensometer were examined and the rock has been classified in RMR, Q System, RSR
and GSI. At Kohala site, the rock surface has been directly classified where the plates
were placed. The sketch showed in Figure 5.19 illustrates the locations where the rock
was classified in the borehole. Ninety (90) data sets thus prepared have been plotted to
develop the correlations.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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152
Figure 5.19: Locations of the rock mass classifications points
Table 5.9 shows the classification rating values at selected locations in four (4) systems.
The effect of blasting may also be observed i.e. for the depth up to 2 m from the rock
surface, the classification rating values in all the systems are on lower side as compared to
the values recorded at 4 and 6 m.
Table 5.9: Classifications of rock where the Deformation Modulus was measured
Test
Location
No.
Side
Depth from
Rock
Surface (m)
RMR Q System RSR GSI
Basha Site
1
Left
0.5 60 13.37 60 48
1.0 69 50.14 68 52
2.0 87 704.86 83 62
4.0 80 252.17 78 58
6.0 88 816.35 84 63
Right
0.5 56 7.43 57 45
1.0 58 9.97 59 46
2.0 80 252.17 78 58
4.0 79 217.73 77 58
6.0 78 188.00 76 57
2 Top
0.5 62 17.94 62 49
1.0 80 252.17 78 58
2.0 78 188.00 76 57
Borehole 76 mm dia
Continued...
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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153
Test
Location
No.
Side
Depth from
Rock
Surface (m)
RMR Q System RSR GSI
4.0 90 1095.02 86 64
6.0 84 453.72 81 60
Bottom
0.5 59 11.55 60 47
1.0 67 37.38 67 51
2.0 65 27.87 65 50
4.0 77 162.32 75 57
6.0 64 24.06 64 50
3
Left
0.5 61 15.49 62 48
1.0 72 77.89 71 54
2.0 68 43.29 67 52
4.0 78 188.00 76 57
6.0 75 121.01 73 56
Right
0.5 52 4.13 54 43
1.0 67 37.38 67 51
2.0 78 188.00 76 57
4.0 84 453.72 81 60
6.0 85 525.48 82 61
4
Top
0.5 66 32.28 66 51
1.0 64 24.06 64 50
2.0 79 217.73 77 58
4.0 83 391.75 80 60
6.0 89 945.48 85 63
Bottom
0.5 65 27.87 65 50
1.0 71 67.26 70 53
2.0 81 292.06 78 59
4.0 78 188.00 76 57
6.0 77 162.32 75 57
5
Left
0.5 64 24.06 64 50
1.0 67 37.38 67 51
2.0 80 252.17 78 58
4.0 86 608.60 83 61
6.0 70 58.07 69 53
Right
0.5 63 20.78 63 49
1.0 70 58.07 69 53
2.0 75 121.01 73 56
4.0 60 13.37 61 48
6.0 90 1095.02 86 64
6 Top
0.5 70 58.07 69 53
1.0 62 17.94 62 49
2.0 74 104.49 72 55
4.0 78 188.00 76 57
6.0 84 453.72 81 60
Bottom 0.5 52 4.13 54 43
Continued...
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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154
Test
Location
No.
Side
Depth from
Rock
Surface (m)
RMR Q System RSR GSI
1.0 65 27.87 65 50
2.0 65 27.87 65 50
4.0 72 77.89 71 54
6.0 79 217.73 77 58
7
Top
0.5 59 11.55 60 47
1.0 60 13.37 61 48
2.0 80 252.17 78 58
4.0 82 338.25 79 59
6.0 90 1095.02 86 64
Bottom
0.5 54 5.54 56 44
1.0 62 17.94 62 49
2.0 63 20.78 63 49
4.0 72 77.89 71 54
6.0 78 188.00 76 57
8
Top
0.5 58 9.97 59 46
1.0 69 50.14 68 52
2.0 75 121.01 73 56
4.0 80 252.17 78 58
6.0 81 292.06 78 59
Bottom
0.5 63 20.78 63 49
1.0 55 6.42 57 45
2.0 58 9.97 59 46
4.0 84 453.72 81 60
6.0 80 252.17 78 58
Kohala Site
1 Left 43 1.10 46 38
Right 48 2.30 51 41
2 Top 45 1.48 48 40
Bottom 36 0.39 41 35
3 Left 45 1.48 48 40
Right 38 0.53 42 36
4 Top 30 0.16 36 31
Bottom 29 0.14 35 31
FJ-1 37 0.46 41 35
FJ-2 32 0.22 37 33
The modulus of deformation have been plotted against corresponding rating values of the
rock in different systems and simple regression analyses have been carried out to develop
correlations among the two parameters. The correlations have been checked for different
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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155
correlation coefficients in linear, logarithmic, power, exponential and polynomial forms
and the best one has been selected. These relations are shown in Figure 5.20 to 5.23.
Figure 5.20: Correlation between modulus of deformation and RMR
Figure 5.21: Correlation between modulus of deformation and Q System
Em = 1.358e0.047RMR
R² = 0.820
0
20
40
60
80
100
120
0 20 40 60 80 100
Em
(GP
a)
RMR
Basha
Kohala
Em = 10.22Q0.324
R² = 0.736
0
20
40
60
80
100
120
0.10 1.00 10.00 100.00 1000.00 10000.00
Em
(GP
a)
Q System
Basha
Kohala
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156
Figure 5.22: Correlation between modulus of deformation and RSR
Figure 5.23: Correlation between modulus of deformation and GSI
Em = 0.756e0.056RSR
R² = 0.836
0
20
40
60
80
100
120
0 20 40 60 80 100
Em
(GP
a)
RSR
Basha
Kohala
Em = 0.343e0.089GSI
R² = 0.840
0
20
40
60
80
100
120
0 10 20 30 40 50 60 70
Em
(GP
a)
GSI
Basha
Kohala
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157
Correlations between rock mass classification systems and deformation modulus is a
traditional tool for estimation of rock mass deformability since Bieniawski introduced the
correlations first time in 1978. The current study is a similar attempt as indicated from
Figures 5.21 to 5.24.
As clear from the above Figures, the general trend of all the graphs is similar. Lower
rating values of classifications in all the systems have yielded lower modulus values and
higher rating values, higher modulus. The coefficients of correlations are in a reasonable
range (from 0.736 to 0.840) and hence good correlations have been obtained.
Consequently, following new correlations have been developed from the current study.
𝐸 = 1.358 𝑒0.047𝑅𝑀𝑅 (5.7)
𝐸 = 10.22 𝑄0.324 (5.8)
𝐸 = 0.756 𝑒0.056𝑅𝑆𝑅 (5.9)
𝐸 = 0.330 𝑒0.074𝐺𝑆𝐼 (5.10)
5.5. VALIDATION BY ARTIFICIAL NEURAL NETWORK ANALYSIS
Artificial Neural Networks (ANNs) are very simplified models similar to human nervous
systems. The models consist of an interconnection of simple processing elements called
neurons, organized in layers. Every neuron of a layer is connected to the neurons in the
subsequent layer and so on.
The information transmission in ANN starts at the first layer where the input data exist. In
this layer, the inputs are weighted and received by all nodes in the next layer. The
weighted inputs are then summed up and a transfer function is applied to produce the
nodal output, which is again weighted and transferred to processing elements in the next
layer. The network adjusts its weights and a learning rule is used until it finds a set of
weights that will produce the input-output mapping which has the smallest possible error.
This process is known as learning or training process.
The most common and simplest type of neural networks consists of three layers: the input
layer, the hidden layer, and the output layer.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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158
In this research, Radial Based Function (RBF) network has been applied for the validation
of correlations. The input layer is responsible for collecting the input information and
formulating the input vector. Hidden nodes of the hidden layer apply nonlinear
transformations to the input vector.
The network input has been selected as RMR, Q system, RSR and GSI rating values in
separate analyses. MATLAB 7.12.0 Neural Network Toolbox has been used for ANN for
validation of deformation modulus. Figure 5.24 shows the M file used as input to generate
modulus values from RMR system.
Figure 5.24: Output file generated from ANN analysis to validate the modulus values
Using the ninety (90) data sets from the Basha and Kohala sites, the modulus values have
been generated in MATLAB. These values have been compared with the values obtained
from regression analysis and found in very close conformity. The comparisons of
modulus measured from in situ tests with the modulus estimated from the current study
(from rock mass classification systems) and estimated from the ANN analysis are shown
in Figures 5.25 to 5.28.
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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159
Figure 5.25: Comparison of modulus of deformation from RMR
Figure 5.26: Comparison of modulus of deformation from Q system
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
imate
d)
GP
a
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(ANN)
± 15%
1:1 Line
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
imate
d)
GP
a
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(ANN)
± 15%± 15%
1:1 Line
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160
Figure 5.27: Comparison of modulus of deformation from RSR
Figure 5.28: Comparison of modulus of deformation from GSI
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
ima
ted
) G
Pa
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(ANN)
± 15%
1:1 Line
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
imate
d)
GP
a
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(ANN)
± 15%
1:1 Line
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161
The Artificial Neural Network analyses revealed that the generated values of deformation
modulus are even better while relating with measured values in in situ tests. It is noted
from above figures that the relative error in all the cases is around ±15% which is
considered to be reasonable keeping in view the numerous factors in such tests.
5.6. COMPARISON OF CORRELATIONS WITH EXISTING CORRELATIONS
During the literature survey, existing correlations between modulus of deformation and
rock mass classification systems have been scrutinized and the most famous and reliable
equations have been selected for comparison with the correlations developed in this study.
The rating values in the rock mass classification system have been put in the relevant
existing equations and thus the modulus obtained are plotted to compare with the current
study in each systems as indicated in Figures 5.29 to 5.32.
Figure 5.29: Comparison with Bieniawski’s equation (from RMR)
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
imate
d)
GP
a
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(Bieniawski)
Em=1.358e0.047RMR (This Study, Eq.5.7)
Em=2RMR-100(Bieniawski, 1978)
± 15%
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
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162
Figure 5.30: Comparison with Barton’s equation (from Q System)
Figure 5.31: Comparison with Sarma’s equation (from RSR)
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
ima
ted
) G
Pa
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(Barton)
E=10.22Q0.324 (This Study, Eq. 5.8)
E=25logQ (Barton, 1993)
± 15%
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
imate
d)
GP
a
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(Sarma)
± 15%
Em=0.756e0.056RSR (This Study, Eq. 5.9)
LogEm=10(RSR-52)/109 (Sarma, 2005)
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
163
Figure 5.32: Comparison with Gokcoeoglu’s equation (from GSI)
It may be observed from the above comparisons that the variance in relative error of
estimation of 95% confidence limit in the correlations developed in this study is better
(within ±15%) than the existing correlations of various researchers (> ±15%). The
equations of Bieniawski (in terms of RMR) and Barton (in terms of Q system) have
slightly greater scatter while the equations of Sarma (in terms of RSR) and Gokcoeglu (in
terms of GSI) have large variations especially for lower RSR and GSI values respectively.
Therefore it is concluded that the newly developed correlations can be presented with
confidence.
5.7. CORRELATIONS OF MODULUS OF ELASTICITY AND MODULUS OF
DEFORMATION
An attempt has been made to correlate the modulus of elasticity or intact modulus (Ei)
obtained from laboratory tests and in situ modulus of deformation. For this purpose
eleven (11) selected samples were taken from the cores extracted from the boreholes
drilled for extensometers in the Basha Adit. These samples were taken to the laboratory
0
10
20
30
40
50
60
70
80
90
100
0 20 40 60 80 100
Em
(Est
ima
ted
) G
Pa
Em (Measured) GPa
E(Measured) vs E(This Study)
E(Measured) vs E(Gokcoeoglu)
± 15%
Em=0.330e0.074GSI (This Study, Eq. 5.10)
Em=0.912e0.0866GSI (Gokcoeoglu, 2003)
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
164
and modulus of elasticity was determined from the slopes of stress-strain curves on 50%
strength.
The modulus of elasticity is actually the measuring of the stiffness of a rock material. It is
defined as the ratio, for small strains, of the rate of change of stress with strain. The
modulus of elasticity of rock materials varies widely with rock type. For extremely hard
and strong rocks, it can be as high as 100 GPa as determined from the hard rock samples
of Basha. The modulus of elasticity and corresponding modulus of deformation obtained
from the Plate Load test at Basha site as described in the preceding pages are given in
Table 5.10.
Table 5.10: Modulus of Elasticity, Modulus of Deformation and Moduli Ratio for
Basha site
Location
No. Side
Distance
from Rock
Surface (m)
Modulus of
Deformation
Em (GPa)
Modulus of
Elasticity
Ei (GPa)
Em/Ei
1 Left 2.00 70.42 98.4 0.716
2 Top 2.00 65.33 114.5 0.571
Bottom 4.00 45.34 82.3 0.551
3 Left 4.00 62.80 102.6 0.612
4 Top 2.00 49.18 89.4 0.550
5 Left 1.00 36.93 78.9 0.468
Right 6.00 63.50 110.5 0.575
6 Bottom 1.00 34.67 80.5 0.431
7 Top 2.00 44.00 83.8 0.525
Bottom 4.00 65.30 125.7 0.519
8 Right 4.00 65.05 123.6 0.526
It is noted that Em is always less than Ei and the average moduli ratio between these two is
found as 0.55. The correlation between Em and Ei is plotted in Figure 5.33. The
correlation has given a high value of regression coefficient i.e. 0.91 for a logarithmic
relation. Another correlation between moduli ratio (Em/Ei) and corresponding RMR value
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
165
has been presented in Figure 5.34 which has been compared with the Bieniawski’s
equation. It is to be noted that in these correlations directional effect of jointing has been
ignored.
Figure 5.33: Correlation between Em and Ei
Figure 5.34: Correlation between moduli ratio and RMR
Em = 63.48ln(Ei) - 237.7
R² = 0.91
0
10
20
30
40
50
60
70
80
90
100
60 70 80 90 100 110 120 130 140
Em
(GP
a)
Ei (GPa)
Em/Ei = 0.242e0.010RMR
R² = 0.548
(This Study)
0.0
0.1
0.2
0.3
0.4
0.5
0.6
0.7
0.8
0 20 40 60 80 100
Em
/Ei
RMR
Em/Ei=e(RMR-100)/36
(Bieniawski et al., 2007)
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
166
As this exercise was done on the samples of core pieces from Basha site, the RMR rating
values are on higher side (>50). The coefficient of correlation is not very high which
indicates a scattered data representation in the graph. However this scatter is always
expected in such attempts due to many uncertainties involved. Consequently, following
correlations between the two kinds of moduli have also been developed in this study.
Both equations are valid for RMR>50.
𝐸𝑚 = 63.48 ln 𝐸𝑖 − 237.7 (5.11)
𝐸𝑚𝐸𝑖 = 0.242𝑒0.01𝑅𝑀𝑅 (5.12)
5.8. SUMMARY
Correlations among modulus of deformation and rock mass classification systems are of
great significance. The current study encompasses the methodology to develop new
correlations between modulus of deformation and four (4) rock mass classification
systems. The moduli of deformation were determined from Plate Load test at Basha and
Kohala sites and Flat Jack tests at Kohala site. From the data of Basha site, the average
values of deformation modulus at each measuring points have been computed and placed
at a single location to observe the variation in deformation modulus with distance from
the rock surface and in this way contours of deformation modulus have been drawn by
using “Surfer” software.
Ninety (90) data sets of deformation modulus and rock mass ratings have been prepared
and plotted to develop the correlations. Two different sites of different quality of rocks
have yielded a wide range of moduli which is very good to develop the correlations. As a
result, four (4) correlations have been developed between deformation modulus and
RMR, Q system, RSR and GSI.
Artificial Neural Network (ANN) analysis has been applied for the validation of
correlations by using the MATLAB software and the correlations were found within
acceptable error limit. Furthermore, the most famous and reliable equations have been
selected from the literature for comparison with the correlations developed in this study.
The comparisons reveal that the variance in relative error of estimation of 95%
confidence limit in the correlations developed in this study is better (within ±15%) than
CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND
VARIOUS ROCK MASS CLASSIFICATION SYSTEMS
167
the existing correlations of various researchers (> ±15%). Modulus of elasticity which is
determined in the laboratory usually, is also correlated with in situ modulus of
deformation with a good coefficient of correlation.
The correlations have been developed for the first time in Pakistan and can be used in
place of existing correlations available in the literature for the rocks of northern area of
Pakistan.
CHAPTER-6
168
CONCLUSIONS AND RECOMMENDATIONS
6.1. INTRODUCTION
Based on the geological data of Basha and Kohala sites, it can be inferred that Basha dam
site mainly comprises of two types of rocks mass namely Gabbronorite (GN) and
Ultramafic Association (UMA). At Kohala Hydropower Project site, two main types of
rock units exist, i.e. Sandstone and Shale. Sandstone is further divided into two types
based on uniaxial compressive strength, i.e. SS-1 and SS-2.
The rocks at Basha and Kohala sites are different in nature and strength, therefore
different rating value ranges in four classification systems were obtained. The rock mass
ratings as determined in all the four systems for the Basha dam site vary in the narrow
range which depicts that rock mass is fairly homogenous, where as for Kohala site, the
rating values have a wide but lower range indicating variable and relatively poor quality
rocks as compared with Basha dam site.
As a part of the research, Plate Load tests and Flat Jack tests performed at Diamer Basha
Dam and Kohala Hydropower Project have been supervised. Data have been analyzed
extensively and used to establish the correlations of Modulus of Deformation with four
rock mass classification systems. Moduli of deformation obtained from Basha site are
higher in range (9.91 to 92.53 GPa) as compared to moduli obtained from Kohala site
(4.32 to 14.57 GPa), again indicating the difference of rock properties at both the sites.
The study encompasses the data of strong rocks of Basha and weak rocks of Kohala;
therefore a wide range has been covered by the developed correlations.
Consequently, along these lines the desired objectives of the research have been achieved
as mentioned in Chapter-1 and new correlations have been proposed which can be used
for the rocks of the northern area of Pakistan.
CHAPTER-6 CONCLUSIONS
169
6.2. CONCLUSIONS
The following conclusions are derived from the research;
Based on the rock mass classifications used in the study, the Basha dam site
mainly comprises of Fair to Very Good quality of Gabbronorite and UMA while
Kohala site consists of Poor to Fair quality of Sandstone / Shale.
In the study, following correlations between various rock mass classification
systems have been developed. These equations have very good regression
coefficients (from 0.835 to 0.901) indicating good correlations between various
systems.
o 𝑅𝑀𝑅 = 6.81 𝑙𝑛𝑄 + 42.34
o 𝑅𝑆𝑅 = 5.92 𝑙𝑛𝑄 + 45.96
o 𝑅𝑆𝑅 = 0.84 𝑅𝑀𝑅 + 10.33
o 𝐺𝑆𝐼 = 0.54 𝑅𝑀𝑅 + 15.43
o 𝐺𝑆𝐼 = 3.69 𝑙𝑛𝑄 + 38.05
Vertical Plate Load Tests have yielded slightly lower values of deformation
modulus as compared to Horizontal Plate Load Tests. Average Em in vertical tests
is 46.1 GPa while for horizontal tests, it is 49.7 GPa.
The modulus of deformation is generally increasing with increase in distance from
the rock surface in all the tests. This increase in modulus is continuous until a
distance is reached after which the modulus is relatively constant with depth. This
distance is about 2 to 3 m from the rock surface for the plate of 0.9 m dia (2 to 3
times the dia of plate).
Following correlations have been developed between modulus of deformation of
rock and rock mass classification systems;
o 𝐸𝑚 = 1.358 𝑒0.047𝑅𝑀𝑅
o 𝐸𝑚 = 10.22 𝑄0.324
o 𝐸𝑚 = 0.756 𝑒0.056𝑅𝑆𝑅
o 𝐸𝑚 = 0.330 𝑒0.074𝐺𝑆𝐼
Using the data sets from the Basha and Kohala sites, the modulus values have
been generated related to 4 classification systems in ANN system. These values
CHAPTER-6 CONCLUSIONS
170
have been compared with the values obtained from regression analysis and found
in very close conformity.
The correlations developed also proved more accurate when their percentage error
was compared with the error produced by the existing correlations.
Correlations between modulus of elasticity of intact rock cores and modulus of
deformation of rock mass have also been developed and the following correlations
have been suggested;
o 𝐸𝑚 = 63.48 ln 𝐸𝑖 − 237.7
o 𝐸𝑚
𝐸𝑖 = 0.242𝑒0.01𝑅𝑀𝑅
6.3. RECOMMENDATIONS FOR FUTURE WORK
The research area can be extended to other sites as well in the same region. It is
recommended to use ANN analysis to construct a model based on the data
obtained in this research and apply to some other sites to compare the measured
and estimated values of modulus of deformation.
It is recommended to explore the effect of variation in Poisson’s ratio on the
modulus of deformation.
The correlations between RQD and modulus of deformation may be established.
Different types of modulus like tangent, recovery and secant may be computed
from the analysis done in the research.
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APPENDIX-A
176
APPENDIX-A
RESULT SHEETS OF LABORATORY TESTS
APPENDIX-A
177
Diamer Basha Dam Project – Typical Result Sheet of Uniaxial Compression Test
Diamer Basha Dam Project – Typical Result Sheet of Indirect Tensile Strength Test
APPENDIX-A
178
Diamer Basha Dam Project – Typical Result Sheet of Point Load Strength Index Test
APPENDIX-A
179
Diamer Basha Dam Project – Typical Result Sheet of Modulus of Elasticity Test
APPENDIX-A
180
APPENDIX-A
181
APPENDIX-A
182
APPENDIX-B
183
APPENDIX-B
DETAILS OF ENGINEERING PROPERTIES TESTS
APPENDIX-B
184
Diamer Basha Dam - Details of Engineering Properties Tests
Sample
No.
Rock Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point Load
Strength
(MPa)
Tensile
Strength
(MPa)
G-35 UMA 60
2.55
G-36 UMA 30
2.19
G-37 UMA
1.29
G-38 UMA
4.85
G-39 UMA
4.70
G-40 UMA 87
9.21
G-41 UMA 62
1.98
G-42 UMA 61
3.16
G-43 UMA 58
4.39
G-44 UMA 46
3.07
G-45 UMA 45
2.63
G-46 UMA 15
2.33
G-47 Gabbronorite 102
6.95
G-48 UMA 34
2.30
G-49 UMA 75
3.92
G-50 Gabbronorite 102
7.67
G-51 Gabbronorite 62
11.85
G-52 UMA 70
7.54
G-53 Gabbronorite 50
2.67
G-54 Gabbronorite 96
5.84
G-55 UMA 31
11.26
G-56 UMA 75
1.71
G-57 UMA 53
2.61
G-58 Gabbronorite 53
5.47
G-59 UMA 96
2.32
G-60 UMA
5.92
G-61 Gabbronorite 158 67.6 0.091 9.21
G-62 Gabbronorite 203 186.0 0.870 9.05
Continued...
APPENDIX-B
185
Sample
No.
Rock Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point Load
Strength
(MPa)
Tensile
Strength
(MPa)
G-63 Gabbronorite 128 38.3 0.260 6.48
G-64 Gabbronorite 177 116.0 0.360 7.85
G-65 Gabbronorite 188 63.4 0.334 11.54
G-66 Gabbronorite 172 93.2 0.182 9.00
G-67 Gabbronorite 103 98.1 0.383 11.20
G-68 Gabbronorite 196 101.0 0.220 11.69
G-69 Gabbronorite 117 59.5 0.083 7.03
G-70 Gabbronorite 116 41.3 0.026 5.23
G-71 Gabbronorite 29 69.2 0.299 7.01
G-72 Gabbronorite 113 91.0 0.324 9.12
G-73 Gabbronorite 129 50.6 0.133 5.68
G-74 Gabbronorite 143 108.0 0.260 7.00
G-75 Gabbronorite 117 49.7 0.131 5.91
G-76 Gabbronorite 92 28.6 0.020 6.45
G-77 Gabbronorite 109 170.0 0.496 7.80
G78 Gabbronorite 158 79.1 0.190 6.82
G-79 Gabbronorite 116 72.8 0.252 6.17
G-80 Gabbronorite 72 30.1 0.029 5.80
G-81 Gabbronorite 75 32.9 0.023 3.34
G-82 Gabbronorite 63 37.8 0.061 3.71
G83 Gabbronorite 168 53.7 0.117 7.54
G-84 Gabbronorite 91 104.0 0.048 4.60
G85 Gabbronorite 87 18.5 0.356 4.14
G-86 Gabbronorite 71 20.4 0.020 3.36
G-87 Gabbronorite 46 11.2 0.022 1.71
G-88 Gabbronorite 69 3.7 0.037 3.10
G-89 Gabbronorite 75 23.7 0.034 3.41
G-90 Gabbronorite 123 - - 7.26 20.39
G-91 Gabbronorite 80 - -. 5.40 8.29
Continued...
APPENDIX-B
186
Sample
No.
Rock Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point Load
Strength
(MPa)
Tensile
Strength
(MPa)
G-92 Gabbronorite 80 68.3 0.088 6.76 15.20
G-93 Gabbronorite 96 161.0 0.630 10.26 16.92
G-94 Gabbronorite 148 228.0 0.952 14.00 14.90
G-95 Gabbronorite 127 37.1 0.042 9.71 13.24
G-96 Gabbronorite 133 30.1 0.017 10.29 11.45
G-97 Gabbronorite 126 37.2 0.033 10.87 12.90
G-98 Gabbronorite 88 45.7 0.222 7.84 11.37
G-99 Gabbronorite 100 86.4 0.209 10.57 12.96
G-100 UMA 125 95.6 0.077 11.06
G-101 UMA 115 154.0 0.238 10.20
G-102 UMA 115 85.8 0.107 8.08
G-103 UMA 138 114.0 0.199 9.36
G-104 UMA 104 55.0 0.022 10.87
G-105 UMA 124 140.0 0.318 10.78
G-106 UMA 126 155.0 0.299 11.64
G-107 UMA 126 340.0 0.187 11.54
G-108 UMA 124 101.0 0.058 12.60
G-109 UMA 78 13.2 0.263 9.03
G-110 UMA 113 173.0 0.354 9.66
G-111 UMA 89 88.3 0.033 8.12
G-112 UMA 80 76.3 0.887 6.35
G-113 UMA 74 36.9 0.025 8.39
G-114 UMA 111 57.8 0.015 10.31
G-115 Gabbronorite 130 47.0 0.130 12.18
G-116 Gabbronorite 52 15.0 0.021 9.52
G-117 Gabbronorite 74 16.5 0.018 6.14
G-118 Gabbronorite 58 13.5 0.280 6.35
G-119 Gabbronorite 99 23.0 0.035 3.39
G-120 Gabbronorite 81 40.3 0.258 5.93
Continued...
APPENDIX-B
187
Sample
No.
Rock Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point Load
Strength
(MPa)
Tensile
Strength
(MPa)
G-121 Gabbronorite 85 43.2 0.804 6.79
G-122 Gabbronorite 108 73.6 0.178 10.78
G-123 Gabbronorite 71 26.3 0.038 7.93
G-124 Gabbronorite 86 129.0 0.076 7.26
G-125 Gabbronorite 89 27.9 0.154 7.93
G-126 Gabbronorite 47 42.4 0.327 5.62
G-127 Gabbronorite 68 55.4 0.232 7.48
G-128 Gabbronorite 158 114.0 0.334 11.79
G-129 Gabbronorite 48 250.0 0.360 10.97
G-130 Gabbronorite 161 233.0 0.205 11.24
G-131 Gabbronorite 32 24.3 0.017 10.31
G-132 Gabbronorite 121 51.3 0.289 7.48
G-133 Gabbronorite 121 45.0 0.060 11.29
G-134 Gabbronorite 74 27.5 0.059 10.47
G-135 Gabbronorite 83 35.8 0.051 9.02
G-136 Gabbronorite 148 56.9 0.169 9.97
G-137 Gabbronorite 92 56.9 0.046 10.95
G-138 Gabbronorite 91 55.1 0.110 10.63
G-139 Gabbronorite 128 146.0 0.639 12.44
APPENDIX-B
188
Kohala Hydropower Project - Details of Engineering Properties Tests
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH2-1 SS-2 21.56
5.00
BH2-2 SS-2
1.39
BH2-3 SS-2 18.71
6.11
BH2-4 SS-2 25.00 33618.41 0.11
BH2-5 SS-2 31.07
BH2-6 SS-2
1.94
BH2-7 SS-2 29.89
15.12
BH2-8 SS-1 75.00 37792.92 0.06 9.44 8.56
BH2-9 SS-2
4.72
BH2-11 SS-2 27.45
6.39
BH15-1 SS-2
1.39
BH15-2 SS-2 53.50
BH15-3B SS-2 48.60
0.83
BH15-3C SS-2 61.00 47065.77 0.21
3.42
BH15-4 SS-2 74.20
3.61
BH15-5 SS-2 35.80
BH15-6 SS-2
6.11
BH15-7 SS-2 13.80
BH15-8 SS-2
3.89 2.79
BH15-11 SS-1 52.00 31633.56 0.09 6.11
BH15-12 SS-1
9.11
BH15-13 SS-2 31.60
5.00
BH15-15 SS-2
1.67
BH8-1 SS-2
5.83
BH8-2 SS-2 39.00 28575.65 0.05
BH8-4 SS-1
10.54
BH10-1 SS-2
4.00
Continued...
APPENDIX-B
189
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH10-5 SS-1 45.57
4.86
BH10-6 SS-1
12.61
BH10-7 SS-1 71.00 23535.19 0.29
BCD1-1 SS-1 115.63
8.33
BCD1-3 SS-1 74.00 61674.06 0.46
0.46
BCD1-4 SS-2 28.60
BCD1-5 SS-2 38.55
3.61
BCD1-6 SS-2 23.00 11600.00 0.05
0.05
BCD1-8 SS-1 32.58
1.67
BCD1-10 Shale 17.60
BCD1-14 Shale 25.72
BCD1-15 Shale 34.00 42178.08 0.15
0.15
BCD1-17 Shale
2.22
BCD1-19 SS-1
2.50
BCD1-20 SS-1 24.00 38615.74 0.32
0.32
BCD2-1 SS-1 42.84
3.33
BCD2-2 SS-1 61.61
BCD2-4 SS-1 31.53
6.25
BCD2-5 SS-2
2.78
BCD2-6 SS-2 22.01
BCD2-7 SS-2 24.00 20314.44 0.10
BCD2-8 SS-2
7.01
BCD2-10 SS-2 28.38
BCD2-12 Shale 31.56
BH4-2 SS-1 21.00 45358.10 0.02
BH4-3 SS-1
6.15 3.99
BH4-4 SS-2
4.75 9.71
BH4-6 SS-2 53.03
Continued...
APPENDIX-B
190
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH4-7 SS-2
3.07
BH4-8 SS-2 81.00 38137.20 0.19
BH4-9 SS-2 25.00 26458.90 0.06 7.54
BH5-1 SS-2 57.06
BH5-2 SS-2
1.96 7.35
BH5-5 SS-1 65.81
3.07
BH6-1 SS-1 39.04
7.65
BH6-2 SS-2
3.07 4.71
BH6-4 SS-2
8.38
BH6-6 SS-2 16.64
7.35
BH6-8 SS-2
8.38
BH13-1 SS-2
1.94
BH13-3 SS-2 14.99
3.24
BH13-4 SS-2
1.39
BH13-5 SS-2
3.82
BH13-6 SS-2 5.19
BH13-7 SS-2
6.94
BH21-1 SS-1 14.82
1.25
BH23-1 Shale 10.16
BH23-2 SS-1 68.60
5.24
BH28-1 SS-2 24.82
5.23
BH28-2 SS-1
11.35
BH31-1 SS-1 13.36
4.25
BH9-1 SS-2 33.15
3.61
BH9-2 SS-2
5.51
BH9-3 SS-2 66.00 38563.60 0.22
BH9-4 SS-2 35.43
Continued...
APPENDIX-B
191
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH9-5 SS-1
4.34
BH9-7 Shale 73.00 31488.30 0.26 5.55
BH9-9 SS-1 92.59
3.05
BH9-10 SS-2 64.00 34672.10 0.41
BH9-11 SS-1
1.94
BH9-14 Shale 39.43
BH9-15 SS-1 61.00 25536.70 0.04 6.39
BH18A-1 SS-1 68.58
2.78
BH18A-2 SS-2 33.72
4.44
BH18A-4 Shale 8.72
BH18A-5 SS-2 30.00 7108.00 0.21 4.16
BH18A-7 SS-1
3.61
BH18A-9 SS-2 58.00 13241.70 0.05 2.50
BH18A-10 SS-1 50.29
BH18A-11 SS-1
4.16
BH18A-12 SS-1 62.87
BH18A-15 SS-1 81.00 15096.50 0.10 4.44
BH18A-16 SS-2 27.06
BH18A-17 SS-2
4.16
BH18A-19 SS-2 17.97
BH18A-20 SS-1 98.61
BH25-1a Shale 50.29
BH25-1b Shale
7.10
BH25-2 SS-1 40.01
BH25-3b SS-2 25.72
BH25-4a SS-1
8.79
BH25-4b SS-1 68.58
BH25-4c SS-1
9.93
Continued...
APPENDIX-B
192
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH26-1 SS-1 37.15
6.66
BH26-2 SS-2 27.43
2.22
BH26-3b Shale 55.34
BH26-4 SS-1
9.93
BH26-5 SS-1
4.77
BH26-6a SS-1
5.00 3.09
BH26-6c SS-1 59.44
7.22
BH26-6d SS-1
10.67
BH26-7b SS-2
1.11
BH26-7d SS-2
3.33
BH26-8 SS-1 13.72
BH26-9 Shale
1.11
BH26-10 SS-2 1.14
1.40
BH26-13a SS-1 44.01
BH26-13b SS-1
6.66
BH26-13c SS-1 40.40
4.77
BH26-14b Shale 16.57
10.00
BH11-1 SS-1 40.00 20234.76 0.24 3.37
BH11-3 SS-2 34.28
6.18 8.36
BH11-6 SS-2
2.53 8.75
BH11-7 SS-2 67.00 24938.52 0.08
BH11-8 SS-2 29.78
BH11-9 SS-2
10.85
BH11-10 SS-1
10.12
BH11-3 SS-1
2.25 7.79
BH11-4 Shale
1.69
BH11-6 SS-1
11.24 13.26
BH11-7 SS-1 90.00 28291.90 0.13
Continued...
APPENDIX-B
193
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH11-8 SS-1 113.51
11.80
BH11-9 SS-1
9.36
BH11-10 SS-1
11.24
BH11-11 SS-1 101.00 52737.50 0.14
BH1-8.0 SS-2 30.55
BH1-1.0 SS-1 42.47
BH1-2.0 SS-2 11.21
1.62
BH1-3.0 SS-1 95.00 31188.50 0.04
9.61
BH1-4.0 SS-2 13.83
BH1-5.0 SS-2
1.64 7.68
BH1-6.0 SS-1 85.00 66910.63 0.49
BH1-10.0 SS-1 42.08
BH1-11.0 SS-2
6.18
BH1-7.0 SS-2
2.48
BH1-12.0 SS-2 14.99
BH1-13.0 SS-2 64.00 69724.27 0.28
5.11
BH1-9.0 SS-2
4.09
BH1-14.0 SS-2
4.68
BH1-17.0 SS-2
3.58
BH1-18.0 SS-2 30.55
4.81
BH1-21.0 SS-1 19.02
BH1-22.0 SS-2 18.00 21614.52 0.50 3.03
BH1-23.0 SS-1 66.00 35756.04 0.12
BH3-1.0 SS-1
4.68
BH3-2.0 SS-2 37.81
BH3-3.0 SS-2
7.77
BH3-4.0 SS-2
2.48
BH3-5.0 SS-2
2.66
Continued...
APPENDIX-B
194
Sample No.
Rock
Type
UCS
(MPa)
Young's
Modulus
(GPa)
Poisson
Ratio
Point
Load
Strength
(MPa)
Tensile
Strength
(MPa)
BH3-7.0 SS-2
3.03
BH3-8.0 SS-2 18.62
BH3-9.0 SS-2 18.00 11970.35 0.27 4.13
BH3-11.0 SS-2
3.29
BH3-12.0 SS-2
1.93
BH3-13.0 Shale
3.50
BH3-14.0 Shale
6.88
BH3-15.0 SS-2 15.00 40695.91 0.59
BH3-16.0 SS-2 25.96
BH3-18.0 SS-1 27.65
3.30
BH14-1.0 SS-1
10.45
BH14-3.0 SS-2
6.33
BH14-4.0 SS-2 101.58
3.09
BH14-5.0 SS-2 44.00 156754.63 0.57
BH14-6.0 SS-2
9.08
BH14-7.0 SS-2
2.68
BH14-8.0 SS-1 27.00 51677.35 0.41 2.48
BH14-9.0 SS-1
4.77
BH14-11.0 SS-2 124.00 79583.12 0.53
BH14-12.0 SS-2 76.18
BH14-13.0 SS-2
5.78 3.27
BH14-14.0 SS-2
7.43
BH14-15.0 SS-2 54.17
BH16-1.0 Shale
1.93
BH16-2.0 Shale 8.07
7.97
BH16-4.0 SS-2
5.61
BH16-5.0 SS-2 34.58
BH16-6.0 SS-2 22.00 17162.55 0.09 3.58 5.52
APPENDIX-C
195
APPENDIX-C
GEOLOGICAL MAPPING OF ADIT 4
DIAMER BASHA DAM
APPENDIX-C
196
Diamer Basha Dam - Typical Geological Mapping (Ch: 75 – 225) of Adit 4
APPENDIX-C
197
APPENDIX-C
198
APPENDIX-C
199
APPENDIX-D
200
APPENDIX-D
BOREHOLE LOGS
KOHALA HYDROPOWER PROJECT
APPENDIX-D
201
Kohala Hydropower Project - Borehole Logs showing Lithology
APPENDIX-D
202
Kohala Hydropower Project - Borehole Logs showing Lithology
APPENDIX-D
203
Kohala Hydropower Project - Borehole Logs showing Lithology
APPENDIX-E
204
APPENDIX-E
PLATE LOAD TEST RESULTS
AT DIAMER BASHA DAM PROJECT
APPENDIX-E
205
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.01.02.03.04.05.06.07.08.09.010.0
0.000 0.200 0.400 0.600
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Right Side Sensor at 0.5 m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Right Side Sensor at 1.0 m)
0.0
1.02.0
3.04.0
5.0
6.07.08.09.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Right Side Sensor at 2.0 m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Right Side Sensor at 4.0 m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.005 0.010
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Right Side Sensor at 6.0 m)
APPENDIX-E
206
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.050 0.100 0.150
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Left Side Sensor at 0.5 m)
0.01.02.03.04.05.06.07.08.09.010.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Left Side Sensor at 1.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.010 0.020 0.030
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Left Side Sensor at 2.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Left Side Sensor at 4.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.002 0.004 0.006
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.1 (Left Side Sensor at 6.0m)
APPENDIX-E
207
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.100 0.200 0.300
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Top Side Sensor at 0.5m)
0.01.02.03.04.05.06.07.08.09.010.0
0.000 0.050 0.100
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Top Side Sensor at 1.0m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Top Side Sensor at 2.0m)
0.01.02.03.04.05.06.07.08.09.010.0
0.000 0.005 0.010 0.015 0.020
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Top Side Sensor at 4.0m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Top Side Sensor at 6.0m)
APPENDIX-E
208
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Bottom Side Sensor at 0.5m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.020 0.040 0.060
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Bottom Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Bottom Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.002 0.004 0.006 0.008 0.010
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Bottom Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.2 (Bottom Side Sensor at 6.0m)
APPENDIX-E
209
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Right Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Right Side Sensor at 1.0m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Right Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Right Side Sensor at 4.0m)
0.01.02.03.04.05.06.07.08.09.010.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Right Side Sensor at 6.0m)
APPENDIX-E
210
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Left Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Left Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Left Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Left Side Sensor at 4.0m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.3 (Left Side Sensor at 6.0m)
APPENDIX-E
211
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.100 0.200 0.300
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Top Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Top Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Top Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Top Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Top Side Sensor at 6.0m)
APPENDIX-E
212
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Bottom Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Bottom Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Bottom Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Bottom Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.002 0.004 0.006
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.4 (Bottom Side Sensor at 6.0m)
APPENDIX-E
213
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Right Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Right Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Right Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Right Side Sensor at 4.0 m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Right Side Sensor at 6.0m)
APPENDIX-E
214
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Left Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Left Side Sensor at 1.0m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.010 0.020 0.030
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Left Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Left Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.5 (Left Side Sensor at 6.0m)
APPENDIX-E
215
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Top Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)Wz (mm)
Test No.6 (Top Side Sensor at 1.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Top Side Sensor at 2.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Top Side Sensor at 4.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.002 0.004 0.006 0.008 0.010
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Top Side Sensor at 6.0m)
APPENDIX-E
216
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Bottom Side Sensor at 0.5m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.050 0.100
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Bottom Side Sensor at 1.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Bottom Side Sensor at 2.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.005 0.010
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Bottom Side Sensor at 4.0m)
0.0
2.0
4.0
6.0
8.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.6 (Bottom Side Sensor at 6.0m)
APPENDIX-E
217
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.100 0.200 0.300
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Top Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Top Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Top Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Top Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Top Side Sensor at 6.0m)
APPENDIX-E
218
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.100 0.200 0.300
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Bottom Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Bottom Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Bottom Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Bottom Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz(mm)
Test No.7 (Bottom Side Sensor at 6.0m)
APPENDIX-E
219
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.050 0.100 0.150 0.200
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Right Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Right Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Right Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Right Side Sensor at 4.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Right Side Sensor at 6.0m)
APPENDIX-E
220
Diamer Basha Dam –Pressure vs Deformation Curves in Plate Load Test
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.100 0.200 0.300
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Left Side Sensor at 0.5m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.020 0.040 0.060 0.080
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Left Side Sensor at 1.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.010 0.020 0.030 0.040
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Left Side Sensor at 2.0m)
0.0
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
0.000 0.005 0.010 0.015
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Left Side Sensor at 4.0m)
0.01.02.03.04.05.06.07.08.09.0
10.0
0.000 0.002 0.004 0.006 0.008
Pre
ssu
re (
MP
a)
Wz (mm)
Test No.8 (Left Side Sensor at 6.0 m)