DEVELOPMENT OF CORRELATION BETWEEN ROCK …

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


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.


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.


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


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).


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


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).


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.


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


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.


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


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


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)


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.


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)


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


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.


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.


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.


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


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...


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.


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.


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.


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


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’


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


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.


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


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


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


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


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


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...


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


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.


40

Figure 2.10: The Geological Strength Index chart (Cai et al., 2004)


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


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


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


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:


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)


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.


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


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.


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.


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


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


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


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


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


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.


58

Figure 3.9: Preparation of samples by cutting the cores

Figure 3.10: Point Load Strength Index test


59

Figure 3.11: Preparations for Modulus of Elasticity test

Figure 3.12: Unconfined Compression Test without strain gauges


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


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


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.


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.


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


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.


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.


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%.


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


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


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


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.


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


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



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



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...


79

Chainage

Ratings for

RMR Rock

Quality PLSI/

UCS RQD

Spacing of

Discontinuities

Condition of


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...


80

Chainage

Ratings for

RMR Rock

Quality PLSI/

UCS RQD

Spacing of

Discontinuities

Condition of


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


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...


82


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...


83


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



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...



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



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...



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



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



89

Figure 4.3: Typical borehole logs of BH 11 and 12 showing lithology at Kohala site



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



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.


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...


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...


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...


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...


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...


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


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...


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...


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...


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...


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


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...



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...



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



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...



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...



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



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.



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



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.



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


Co

rrea

ltio

n C

oef

fici

ent

0

0.2

0.4

0.6

0.8

1


Corr

ealt

ion C

oef

fici

ent

0

0.2

0.4

0.6

0.8

1


Corr

ealt

ion C

oef

fici

ent

0

0.2

0.4

0.6

0.8

1


Corr

ealt

ion C

oef

fici

ent



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



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



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



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



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.



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)



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)



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



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

CHAPTER-5 CORRELATIONS BETWEEN DEFORMATION MODULUS AND


123

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



124

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


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


6 Right X-Cut Vertical 18.0 3.85 Dip > 45

o, Joint striking


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


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



125

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...



126

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.



127

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.



128

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



129

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).



130

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.



131

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



132

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



133

Figure 5.6: Flat jack and hydraulic pump

Figure 5.7: Installation of plates, Flat Jacks and spacers



134

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.



135

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)



136

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


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...



137


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


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


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


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...



138


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


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


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



139

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



140

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

)


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

)


Test No. 3 (Horizontal)

Left

Right

0

20

40

60

80

100

0 1 2 3 4 5 6 7

Em

(G

Pa

)


Test No. 4 (Vertical)

Top

Bottom

0

20

40

60

80

100

0 1 2 3 4 5 6 7

Em

(G

Pa

)



Left

Right

0

20

40

60

80

100

0 1 2 3 4 5 6 7

Em

(G

Pa

)



Top

Bottom

0

20

40

60

80

0 1 2 3 4 5 6 7

Em

(G

Pa

)



Top

Bottom

0

20

40

60

80

100

0 1 2 3 4 5 6 7

Em

(G

Pa

)



Left

Right



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)


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



142

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 (%)



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.



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



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.



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)



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.



148

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



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)



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



151

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.



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)


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...



153

Test

Location

No.

Side

Depth from

Rock

Surface (m)


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...



154

Test

Location

No.

Side

Depth from

Rock

Surface (m)


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



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



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



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.



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.



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



± 15%± 15%

1:1 Line



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



± 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



± 15%

1:1 Line



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(Bieniawski)

Em=1.358e0.047RMR (This Study, Eq.5.7)

Em=2RMR-100(Bieniawski, 1978)

± 15%



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(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(Sarma)

± 15%

Em=0.756e0.056RSR (This Study, Eq. 5.9)

LogEm=10(RSR-52)/109 (Sarma, 2005)



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(Gokcoeoglu)

± 15%

Em=0.330e0.074GSI (This Study, Eq. 5.10)

Em=0.912e0.0866GSI (Gokcoeoglu, 2003)



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



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)



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



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.

REFERENCES

171

REFERENCES

ASTM D4394-84 (1998), “Standard Test Method for Determining the In-situ Modulus of

Deformation of Rock Mass Using the Rigid Plate Loading Method” American Society

for Testing and Materials International, USA.

ASTM D4395-84 (1998), “Standard Test Method for Determining the In-situ Modulus of

Deformation of Rock Mass Using the Flexible Plate Loading Method” American

Society for Testing and Materials International, USA.

ASTM D4403-84 (2000), “Standard Practice for Extensometers Used in Rock” American

Society for Testing and Materials International, USA.

ASTM D4729-04 (2004), “Standard Test Method for In-situ Stress and Modulus of

Deformation Using the Flat Jack Method”, American Society for Testing and Materials

International, USA.

ASTM D5878-08, (2008), “Standard guides for Using Rock Mass Classification Systems

for Engineering Purposes”, American Society for Testing and Materials International,

USA.

Barton, N., Lien, R., & Lunde, J., (1974), “Engineering Classification of Jointed Rock

Masses for the Design of Tunnel Support,” Rock Mechanics, Vol. 6, pp. 189-236.

Barton, N., (2000), “TBM Tunnelling in Jointed and Faulted Rock”, Balkema,

Rotterdam, pp. 173.

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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


7.67


11.85

G-52 UMA 70

7.54


2.67


5.84

G-55 UMA 31

11.26

G-56 UMA 75

1.71

G-57 UMA 53

2.61


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


APPENDIX-D

203


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

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APPENDIX-E

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APPENDIX-E

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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)


APPENDIX-E

220


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)


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)


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)


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)


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)