Engineering Geological and Geotechnical Study of Upper Trisuli - 3a Hydroelectric Project, Nuwakot...

ENGINEERING GEOLOGICAL AND GEOTECHNICAL

STUDY OF UPPER TRISULI - 3A HYDROELECTRIC

PROJECT, NUWAKOT AND RASUWA DISTRICT,

CENTRAL NEPAL

A DISSERTATION

(COURSE NO: GEO 639)

SUBMITTED BY

NARAYAN KRISHNA GANESH

EXAM ROLL NO: 650 (2008)

SUBMITTED TO

CENTRAL DEPARTMENT OF GEOLOGY

INSTITUTE OF SCIENCE AND TECHNOLOGY

TRIBHUVAN UNIVERSITY

KIRTIPUR, KATHMANDU

NEPAL

IN PARTIAL FULFILLMENT OF THE REQUIREMENT FOR THE MASTER’S

DEGREE OF SCIENCE IN GEOLOGY

(ENGINEERING GEOLOGICAL TECHNIQUES)

2010

i

It is certified that Mr. Narayan Krishna Ganesh has worked satisfactorily for his Master’s

Degree Dissertation under my guidance and supervision. He has worked enthusiastically

with sincere interest. The dissertation entitled “ENGINEERING GEOLOGICAL AND

GEOTECHNICAL STUDY OF THE UPPER TRUSULI - 3A HYDROELECTRIC

PROJECT, NUWAKOT AND RASUWA DISTRICT, CENTRAL NEPAL” embodies

the candidate’s own work and will be helpful to assess the geotechnical design of

underground structures of the project. I, hereby, recommend the dissertation for approval.

…………………………………

Supervisor

Mr. Jayandra Man Tamrakar

Manager (Senior Geologist)

Engineering Service

Nepal Electricity Authority

ii

It is my pleasure to approve the Dissertation entitled “ENGINEERING GEOLOGICAL

AND GEOTECHNICAL STUDY OF THE UPPER TRISULI - 3A HYDROELECTRIC

PROJECT, NUWAKOT AND RASUWA DISTRICT, CENTRAL NEPAL”

accomplished by Mr. Narayan Krishna Ganesh. The work represents entirely his

individual research work with technical assistance from the Central Department of

Geology and Nepal Electricity Authority.

I recommend the dissertation for approval.

…………………………………….

Internal Supervisor

Dr. Kamala Kant Acharya

Lecturer

Central Department of Geology

iii

The M.Sc. Dissertation entitled “ENGINEERING GEOLOGICAL AND

GEOTECHNICAL STUDY OF THE UPPER TRISULI - 3A HYDROELECTRIC

PROJECT, NUWAKOT AND RASUWA DISTRICT, CENTRAL NEPAL” was

submitted and successfully presented by Mr. Narayan Krishna Ganesh to the Central

Department of Geology, Tribhuvan University, Kirtipur, Kathmandu, Nepal. We, hereby,

certify that this work fulfill the partial requirement for obtaining the Master’s Degree of

Science in Geology.

…………………………………….

Dr. Suresh Das Shrestha

Acting Head


…………………………………….

External Examiner

Mr. Churna Bahadur Oli

Senior Engineering Geologist

Department of Electricity Development

…………………………………….

Supervisor

Mr. Jayandra Man Tamrakar

Manager (Senior Geologist)

Nepal Electricity Authority

…………………………………….

Internal Supervisor

Dr. Kamala Kant Acharya

Lecturer


iv

ACKNOWLEDGEMENT

First, I express my gratitude to my supervisor, Mr. Jayandra Man Tamrakar, Nepal

Electricity Authority, for his continuous support in my dissertation. He showed me

different ways to approach a problem and the need to be persistent to accomplish any

goal. I also thank my co-supervisor, Dr. Kamalakant Acharya, Tribhuvan University, with

whom I explored the ideas, organization, and requirements and complete the writing of

this dissertation. Without their encouragement and constant guidance, I could not have

finished this dissertation.

Besides my supervisors, I would like to thank my teachers Prof. Dr. Megh Raj Dhital and

Prof. Dr. Prakash Chandra Adhikary, Tribhuvan University, who gave insightful

comments, reviewed my work and encouraged to be a better researcher.

A special thanks goes to Nepal Electricity Authority for providing all necessary data and

information regarding the project without whom my study would have been lame and

incomplete. I am also grateful to Mr. Sunil Shrestha, Nepal Electricity Authority and Mr.

Ujjwal Raghubanshi for providing necessary documents and valuable suggestions.

I would like to thank my friends Mr. Navin Shakya, who assisted in my fieldwork, Mr.

Binod Maharjan and Mr. Sudip Shrestha, who helped me with CAD, GIS and other

computer related works, and proofread and mark up my papers and chapters, and Mr.

Bishnu Thapa, who helped me in laboratory works. I also extend my special thanks to all

my friends, teachers, members and staffs of Central Department of Geology for helping

me on various stages of my work.

Last but not the least, I thank my parents and other family members for their

unconditional support and encouragement to pursue my interest. All that I have achieved

is because of you.

Mr. Narayan Krishna Ganesh

2010

v

“ENGINEERING GEOLOGICAL AND GEOTECHNICAL STUDY OF THE

UPPER TRISULI - 3A HYDROELECTRIC PROJECT, NUWAKOT AND

RASUWA DISTRICT, CENTRAL NEPAL”

ABSTRACT

Narayan Krishna Ganesh

Central Department of Geology, Tribhuvan University, Kritipur

The Upper Trisuli - 3A Hydroelectric Project is a running type hydroelectric project, located in

Rasuwa and Nuwakot districts, Bagmati Zone, Central Nepal. It is of capacity 60 MW with

discharge of 51 cumecs and head of 144.5 m. The study area lies with in the latitude 28º 01’ 23“ N

to 28º 04’ 04“ N and loongitude 85º 10’ 47“ E to 85º 12’39“ E. The study includes the geological,

engineering geological and geotechnical study of the Upper Trisuli - 3A Hydroelectric Project.

Geologically, the study area lies in the Kuncha Group of the Lesser Himalaya Metasediments of

the Central Nepal. The Main Boundary Thrust (MBT) is located at about 80 km south of project

area and Main Central Thrust (MCT) about 25 km north of the project area. A local anticline is

expected to exist in the project area whose axis lies in the tunnel alignment. The rock of the study

area is represented by Schist Unit and Gneiss Unit. In general, strike of foliation in project area

varies from NNE-SSW to NNW-SSE dipping NE-SW.

The headwork site lies in the Schist Unit and it is geologically suitable for construction of

hydraulic structures. The dam and desanding basin will be founded on the alluvium deposits. The

headrace tunnel is horse shoe shaped 4142 m long with diameter of 6.2 m and oriented NNE to

SSW. Only 5% of the headrace tunnel will pass through Schist Unit and remaining 95% will pass

through the Gneiss Unit. The rock along the headrace tunnel are poor to good with RMR and Q

values 25-68 and 3.75-13.69 respectively. The foliation is almost perpendicular to tunnel

orientation throughout which is favorable condition for tunnel excavation if drive with dip. The

surge shaft and powerhouse are underground and consist of gneiss of fair quality.

Average in situ deformation modulus (Em) ranges between 9.28 and 30.20 GPa. Vertical stress

(σv) and horizontal stress (σh) as well as horizontal to vertical stress ratio (k) along the

underground structures ranges 2.024 – 8.100 MPa, 2.966 – 8.511 MPa and 0.771 – 2.493

respectively. Damage index (Di) along underground structure ranges 0.049 – 0.195.

Support design for construction of the headrace tunnel based on different system suggests the

combination of local to systematic bolting and reinforced shotcrete as per requirement.

vi

TABLE OF CONTENT

ACKNOWLEDGEMENT ................................................................................................. iv

ABSTRACT ....................................................................................................................... v

LIST OF FIGURES ........................................................................................................... ix

LIST OF TABLES ........................................................................................................... xii

ACRONYMS ................................................................................................................. xiii

SYMBOLS ...................................................................................................................... xiv

CHAPTER ONE ................................................................................................................. 1

INTRODUCTION ........................................................................................................... 1

1.1 LOCATION AND ACCESSIBILITY ................................................................... 2

1.2 TOPOGRAPHY AND DRAINAGE ..................................................................... 2

1.3 VEGETATION ...................................................................................................... 3

1.4 SOCIO-ECONOMIC CONDITION ...................................................................... 4

1.4 PROJECT IN GENERAL ...................................................................................... 4

CHAPTER TWO ................................................................................................................ 7

OBJECTIVE AND METHODOLOGY .......................................................................... 7

2.1 OBJECTIVE .......................................................................................................... 7

2.2 METHODOLOGY ................................................................................................ 7

2.2.1 Desk study ....................................................................................................... 8

2.2.2 Field study ....................................................................................................... 8

2.2.3 Data processing, Interpretation and Report writing ........................................ 9

CHAPTER THREE .......................................................................................................... 10

GEOLOGY OF THE STUDY AREA ........................................................................... 10

3.1 GEOLOGY OF THE NEPAL HIMALAYA. ...................................................... 10

3.1.1 Terai Plain ..................................................................................................... 10

vii

3.1.2 Sub-Himalaya (Siwalik) ................................................................................ 11

3.1.3 Lesser Himalaya ............................................................................................ 12

3.1.4 Higher Himalaya ........................................................................................... 13

3.1.5 Tibetan Tethys Himalaya .............................................................................. 13

3.2 REVIEW OF PREVIOUS GEOLOGICAL WORK IN CENTRAL NEPAL ..... 14

3.3 GEOLOLOGY OF THE CENTRAL NEPAL ..................................................... 17

3.3.1 Chautara-Okhaldunga Metasediment Zone .................................................. 18

3.3.2 Kathmandu Nappe ........................................................................................ 19

3.3.3 Grokha-Nawakot Metasediment Zone .......................................................... 20

3.4 GEOLOGY OF THE PROJECT AREA.............................................................. 21

3.4.1 Schist Unit ..................................................................................................... 21

3.4.2 Gneiss Unit.................................................................................................... 22

CHAPTER FOUR ............................................................................................................ 25

SEISMICITY OF THE PROJECT AREA .................................................................... 25

4.1 SEISMICITY OF NEPAL ................................................................................... 25

4.2 SEISMICITY EVALUATION ............................................................................ 25

4.3 NEPALESE STANDARD ................................................................................... 26

4.4 INDIAN STANDARD ......................................................................................... 27

CHAPTER FIVE .............................................................................................................. 31

ENGINEERING GEOLOGICAL INVESTIGATION OF THE PROJECT AREA ..... 31

5.1 ENGINEERING GEOLOGICAL CONDITION OFTHE HEADWORKS ........ 31

5.1.1 Diversion Weir .............................................................................................. 32

5.1.2 Intake Canal .................................................................................................. 37

5.1.3 Aquaduct ....................................................................................................... 37

5.1.4 Desanding Basin ........................................................................................... 37

viii

5.2 ENGINEERING GEOLOGICAL CONDITION OF DIFFERENT

STRUCTURES .......................................................................................................... 38

5.2.1 Intake Portal .................................................................................................. 38

5.3 Engineering Geological Condition of Headrace Tunnel .................................. 40

5.4 Engineering Geological Condition of Surge Tank ........................................... 47

5.5 Engineering Geological Condition of Inclined Shaft And Penstock Tunnel ... 50

5.6 Engineering Geological Condition of Powerhouse Site .................................. 50

5.7 Engineering Geological Condition Of Tailrace Tunnel ................................... 54

CHAPTER SIX ................................................................................................................ 55

GEOTECHNICAL STUDY OF THE UNDERGROUND STRUCTURES ................. 55

6.1 STRESS ANALYSIS ALONG UNDERGROUND STRUCTURES ................. 56

6.1.1 Estimation of In situ Deformation Modulus ................................................. 56

6.1.2 In situ Stress Analysis ................................................................................... 57

6.1.3 Determination of Elastic and Plastic Behavior of Rock ............................... 58

6.1.4 Determination of Rock Mass Strength along the Headrace Tunnel. ............ 60

6.2 UNDERGROUND WEDGE STABILITY ANALYSIS ..................................... 65

6.3 ROCK SUPPORT DESIGN ................................................................................ 79

6.3.1 Rock Support Design Based On Rock Quality Designation (RQD) ............. 80

6.3.2 Rock Support Design Based on Rock Mass Rating (RMR) ......................... 81

6.3.3 Rock Support Design Based On Tunneling Quality Index (Q) .................... 82

6.3.4 Rock Support Design based on Empirical Design Recommendation

According to U.S. Corps of Engineers................................................................... 86

CHAPTER SEVEN .......................................................................................................... 90

CONCLUSIONS............................................................................................................ 90

7.1 CONCLUSIONS.................................................................................................. 90

REFERENCES .............................................................................................................. 93

ANNEXS ....................................................................................................................... 98

ix

LIST OF FIGURES Figure 1.1: Location map of the study area. ....................................................................... 3

Figure 1.2: Drainage map of the study area ..................... Error! Bookmark not defined.

Figure 3.1: Generalized geological map of Himalaya (Ganser, 1964) ............................. 11

Figure 3.2: Geological map of the Central Nepal Himalaya after Colchen et al., 1986,

Modified by Rai, 2001 .................................................................................... 18

Figure 3.3: Geological map of the study area with cross section along AB. ............Error!

Bookmark not defined.

Figure 4.1: Micro seismicity epicenter map of Nepal (prepared by National

Seismological Center/Department of Mine and Geology). ............................ 27

Figure 4.2: Seismic hazard map of Nepal (published by Department of Mines and

Geology) ......................................................................................................... 28

Figure 4.3: Seismic risk map of Nepal (source: Kaila K.I., Gaur U.K. and Narain, H

(1972)). ........................................................................................................... 29

Figure 4.4: Seismic hazard map of India (Source: Kalia, k.l, Gaur U.K. and Narain, H

(1972)) ............................................................................................................ 30

Figure 5.1: (a) Upstream view and (b) Down stream view of the headwork site............. 32

Figure 5.2: Detail engineering geological map of headwork site.Error! Bookmark not

defined.

Figure 5.3: Geological section along the weir axis .......... Error! Bookmark not defined.

Figure 5.4: Stereographic projection of discontinuities measured around Diversion Weir

........................................................................................................................ 36

Figure 5.5: Photograph of Intake Portal facing 250º ........................................................ 39

Figure 5.6: Stereographic projection of discontinuities measured at Intake Portal .......... 39

Figure 5.7: Engineering geological map of headrace tunnel.Error! Bookmark not

defined.

x

Figure 5.8: Geological cross section along headrace tunnelError! Bookmark not

defined.

Figure 5.9: Stereographic projection of joints measured along road cut section at

Chepleti. ......................................................................................................... 44

Figure 5.10: Stereographic projection of discontinuities along road cut section at

downhill of Katunje guan. .............................................................................. 45

Figure 5.11: Stereographic projection of discontinuities measured along road cut section

downhill of Diyale guan. ................................................................................ 46


at down hill of Danda guan. ........................................................................... 47

Figure 5.13: Geological cross section from surge tank to tailrace tunnel. ................Error!


Figure 5.14: Detailed engineering geological map of powerhouse area. ..................Error!


Figure 5.15: Photograph of powerhouse site near Paire guan. ......................................... 53


near Simle. ...................................................................................................... 53

Figure 6.1: Stereoplot of major joint sets within Ch 0+000 m to 0+876 m ..................... 65

Figure 6.2: Wedges expected in tunnel in between Ch 0+000 m to 0+876 m ................. 66

Figure 6.3: Support applied to stabilize Wedge No 6 ...................................................... 67


Figure 6.5: Stereoplot of major joint set within Ch 0+876 m to Ch 1+938 m ................. 68

Figure 6.6: Wedges expected in the tunnel within Ch 0+876 m to Ch 1+938 m ............. 68

Figure 6.7: Support applied to stabilize Wedge No 6 and 8 ............................................. 69


Figure 6.9: Stereoplot of major joints within Ch 1+938 m to 2+600 m ........................... 71

Figure 6.10: Wedges expected in tunnel within Ch 1+938 m to 2+600 m....................... 71

Figure 6.11: Support applied to stabilize Wedge No 7 and 8 ........................................... 72

xi

Figure 6.12: Support applied for Wedge No 5 ................................................................. 73

Figure 6.13: Stereplot of major joint sets within Ch 2+600 m to Ch 4+142 m ................ 73

Figure 6.14: Wedges expected in tunnel within Ch 2+600 m to Ch 4+142 m ................. 74



Figure 6.17: Stereoplot of major joint set in powerhouse cavern .................................... 76

Figure 6.18: Wedge expected in powerhouse cavern. ...................................................... 76


Figure 6.20: Support applied for Wedge No 7 and 8 ....................................................... 78

Figure 6.21: Estimated support categories based on the tunneling quality index (Q) ...... 85

xii

LIST OF TABLES Table 3.1: Stratigraphy of Kathmandu Complex (Stocklin and Bhattarai, 1977; Stocklin,

1980) ............................................................................................................... 20

Table 5.1: Rock mass classification along headrace tunnel based on RMR value........... 48

Table 5.2: Rock mass classification along headrace tunnel based on Q value ................ 48

Table 6.1: Estimation of in situ deformation of rock along underground structures ....... 57

Table 6.2: Estimation of in situ horizontal and vertical stress along underground

structures ........................................................................................................ 59

Table 6.3: Damage index of rock mass along underground structures. ........................... 60

Table 6.4: Determination of rock mass strength parameter, mb and s .............................. 62

Table 6.5: Values of mi for intact rock (Marrinos and Hoek, 2001) ................................ 63

Table 6.6: Analysis of rock strength using Roclab. ......................................................... 64

Table 6.7: Support recommendations for Tunnels in Rock (6 m to 12 m dia.) based on

RQD (after Deere et al. 1970) ........................................................................ 80

Table 6.8: Estimation of rock support for underground structures based on RQD. ......... 81

Table 6.9: Geomechanics classification guide for excavation and support in rock tunnels

after Bieniawski (1989) .................................................................................. 82

Table 6.10: Estimation of excavation and support in underground structures based on

RMR. .............................................................................................................. 83

Table 6.11: Value of the ESR for the different types of the excavation category (Barton

et al. 1974) ...................................................................................................... 84

Table 6.12: Estimation of rock support based on Q ......................................................... 86

Table 6.13: Recommendation of rock bolt reinforcement based on empirical design

recommendation to U.S. corps of engineers. .................................................. 87

Table 6.14: Estimation of rock support based on empirical design recommendation to

U.S. corps of engineers ................................................................................... 88

Table 6.15 Rock support for Upper Trisuli – 3A HEP. .................................................... 89

xiii

ACRONYMS

DMG : Department of Mine and Geology

ESR : Excavation Support Ratio

GSI : Geological Strength Index

HEP : Hydro Electric Project

HFT : Himalayan Frontal Thrust

HHCs : Higher Himalayan Crystalline System

MBT : Main Boundary Thrust

MCT : Main Central Thrust

MPa : Mega Pascal

MT : Mahabharat Thrust

MW : Mega watt

NEA : Nepal Electricity Authority

Q : Rock Tunneling Quality Index

RMR : Rock Mass Rating

RMR89 : Rock Mass Rating proposed by Bieniawski

(1989)

RQD : Rock Quality Designation

STDS : South Tibetan Detachment System

UCS : Uniaxial Compressive Strength

xiv

SYMBOLS

σ1 : Major Principal Stress

σ2 : Intermediate Principal Stress

σ3 : Minor Principal Stress

Em : In situ Deformation Modulus

σv : Vertical Stress

σh : Horizontal Stress

k : Ratio of Horizontal to Vertical stress

γ : Unit Weight

z : Depth Below Surface

Di : Damage Index

σmax : Maximum Tangential Boundary Stress

σc : Unconfined Compressive Strength

σcs : Uniaxail Compressive Strength

mb and s : Material Constant

σt : Uniaxial Tensile Strength

B : Tunnel Width

De : Equivalent Dimension

L : Rock Bolt Length

Jn : Joint Set Number

Jr : Joint Roughness Number

CHAPTER ONE

INTRODUCTION Nepal, by virtue of its natural setting with great variation in altitude from the Himalayas

to the Lowlands of the Terai over a relatively narrow width (about 200 km) combined

with abundant snowmelt and monsoon water offers tremendous energy potential for

generating hydropower. The major river basins of Nepal are the Koshi, the Gandaki, the

Karnali and the Mahakali, which originate in the high Himalaya or the Tibetan Plateau

and have varying proportion of snow contribution in their flow. The West Rapti, the

Babai, the Kamala are the other rivers originating from the Mahabharat range with little

or no snow contribution. Hydropower is one of the major natural resources of the

economic development of our country. The gross theoretical-potential of Nepal’s rivers

based on average flows has been estimated about 83,000 mega watt (MW). Only about

51% (42,000 MW) of theoretical potential is economically feasible.

According to Nepal Electricity Authority, the total installed capacity including private

and other sector is 617.380 MW comprising 563.870 MW of hydroelectric power, 53.410

MW of diesel power plants and 0.1 MW of solar power plant (NEA, 2007/08). In spite

of abundance potentiality of hydropower, very small amount has been exploited till now

and Nepal has been facing power crisis since many years. Governmental and many

private sector parties are in the field of investigation, implementation and promotion of

hydro-project. For the construction and implementation of hydropower projects, detailed

geological and engineering geological investigations are of prime importance, which help

to construct the safe, efficient and cost effective infrastructures. Geologically, Nepal lies

in the tectonically active zone of the world as well as in the most seismically active zone.

Steep slopes, prevalence of fragile geology, concentrated precipitation and flood, high

river gradient and an alarming rate of deforestation play an important roleto mass

movement. Considering all these natural processes, a complete geological study is quite

essential for the construction of hydropower in Nepal.

Engineering geological and geotechnical investigation of the Upper Trisuli - 3A

Hydroelectric Project (HEP) has been carried out for the partial fulfillment of the

requirement for the master’s degree of science in geology. The major part of the study is

confined to the area where proposed major hydraulic structures such as headworks,

2

headrace tunnel, powerhouse and tailrace tunnel sites. The intake site lies in the schist and

the power house site lies in the gneiss of the Ulleri Formation. During the study,

geological map, engineering geological map and their respective cross sections were

prepared. The area has been studied especially in the engineering geological and

geotechnical aspect. The dissertation is the outcome of overall two weeks of fieldwork

and about twenty four weeks of table work.

1.1 LOCATION AND ACCESSIBILITY

Upper Trisuli - 3A Hydroelectric Project is located in Rasuwa and Nuwakot districts,

Bagmati Zone, Central Nepal. The project site is about 80 km northwest of Kathmandu.

The study area lies between latitude 28º 01’ 23“ N to 28º 04’ 04“ N and longitude 85º 10’

47“ E to 85º 12’ 39“ E (Figure 1.1) on topographical map of scale 1:50,000 (Sheet No.

2885 13, Department of Survey/Government of Nepal, 1996). Geologically, the study

area lies in the Lesser Himalaya of the Central Nepal. The study area covers

approximately 40 km2.

The study area is accessible by gravelly motorable road and take about one and half hour

from Trisuli Bazar which is linked to Kathmandu by Kathmandu-Dhunche road. The

villages around the study area are accessible by the foot trails. Due to damage of

suspension bridges, the left bank of the study area is little difficult for access.

1.2 TOPOGRAPHY AND DRAINAGE

The study area consist of varied topography. Ridges, saddles, spurs, alluvial fans and

flood plains are main topographic features. The area has number of streams and gullies

which shows rugged topography. The steep rocky cliffs and moderate soil terraces present

in the area indicate that the surface relief is strongly controlled by the lithological

variation.

The area shows varying altitudes ranging from 720 m to 1866 m from mean sea level.

The lowest elevation is at the confluence of the Andheri Khola and the Triuli River at the

southern part of project area, about 500 m donwnstream along the Trisuli River from the

proposed powerhouse site. The highest elevation is at north of Mailung Dobhan.

3

The Trisuli River is the main river of the area which flow from NNE to SSW within

project area. This perennial river is snow fed and is originated from the Ganesh Himal.

The Mailung Khola and the Andheri Khola are the main tributaries of the Trisuli River

within the project area. The Mailung Khola joins the Trisuli River at Mailung Dhobhan,

northern part of project area about 1 km upstream from proposed dam site. The Andheri

Khola meets the Trisuli River near Shanti Bazar at Archale. The confluence is about

500m downstream from the proposed powerhouse site. Besides, there are many other

seasonal tributaries most of which make water fall to join Trisuli River but they have not

cut bedrock incisely. The overall drainage pattern of the project area is sub parallel

(Figure 1.2).

Figure 1.1: Location map of the study area.

1.3 VEGETATION

The study area has sub-tropical and deciduous trees. The hills are covered with the forest.

Dense mixed forests to sparsely vegetal are found in the area. The gentle slopes and flat

land is cultivated. The Sallo (Pinus roxburghii), Sal (Shorea robusta), Chilaune (Schirna

wallichi), Uttis (Alnus nepalenstis) are dominant species on the forest.

4

1.4 SOCIO-ECONOMIC CONDITION

Agriculture is the main occupation of the people in this area. The main food crops grown

here are rice, wheat and maize. The geographical setting of the area prevents to being

irrigated properly. The river terraces, gentle slope and ridges of the hills are used for

cultivation. The local hotels and shops are also the small scale business for a few peoples.

Some are engaged in governmental service. The economic condition in Shanti Bazar and

Dandagaun area is quite good but rest of the area are economically weak. Most of the

inhabitants belong to Brahman, Chhetri and Magar. The houses are built by locally

available building materials like stone and woods.

1.4 PROJECT IN GENERAL

i. Type of project Run of River Hydropower

ii. Hydrology

Name of river Trusuli

Reference hydrology Betrawati St. no. 447

Catchment area 4542 sq. km

Design discharge 51 cumecs based on 70% exceedance flow

iii. Geology

Regional geology Lesser Himalaya

Geology of the project area Gneiss, Schist

iv. Project General Description

Gross head 144.5 m

Type of headworks Gated weir with side intake

Design flood 2424 cumecs based on 1:1000 year flood

Full supply level El: 870.5 m from msl

Undersluice gate size 4 nos. (11.6 m × 10 m)

5

Intake type Side intake

Intake channel length 148 m

Desander Twin Berri type

Desander size 95 m × 30 m × 9.2 m (L× B× H)

Headrace tunnel length 4142 m

Headrace tunnel shape D type (excavated) and circular (finished)

Headrace tunnel size 5.4 m for concrete lined and 5.9 m for

shotcrete

Shotcrete lined portion about 60% of total length

Surge shaft Restricted orifice type 17 m dia. 37.7 m high

Inclined shaft Length 168.27 m, diameter 4 m

Pressure tunnel Length 86.6 m, diameter 4.0 m to 2.0 m

Powerhouse type Underground

Powerhouse size 42.6 m × 14 m × 30.2 m

Turbine type Vertical Francis

Installed capacity 60 MW (2 × 30 MW)

Switchyard size 2 nos of 50 m × 15 m

Tailrace conduit D type 6.2 m × 5.02 m size, 115 m length

and twin conduits 25 m length

Tail water level El. 726 m

v. Power Generation

Minimum power generation 43.75 MW

Annual average energy 489.76 GWh (gross)

7

CHAPTER TWO

OBJECTIVE AND METHODOLOGY

The present study is concerned with geological, engineering geological and geotechnical

investigation of the Upper Trisuli - 3A Hydroelectric Project (HEP) for the purpose of

M.Sc. dissertation which is solely an academic.

2.1 OBJECTIVE

The main objective of the present study is to collect geological, engineering geological

and geotechnical information in order to assess the technical prospect of the Upper Trisuli

- 3A HEP. The objectives of the study are summarized below.

1. To study the geology of the area and to prepare the geological map on the scale of

1:50000 with its cross section.

2. To study the engineering geological condition of the project area and to prepare

detailed engineering geological map of intake site and powerhouse site at 1:1000

scale and that of tunnel alignment area at 1:15000 scale.

3. Classification of the rocks of study area on the basis of Rock Mass Rating (RMR)

and Q-Value system.

4. To study the rock mass condition of the tunnel alignment and to classify them for

the design of support pattern.

5. To carry out the stereographic analysis of joints and their interpretation.

6. To carry out the preliminary study to find out the in situ-stress condition along

underground structures by using geotechnical parameters.

7. To design the support pattern for the underground structures on the basis of Rock

Mass Rating (RMR) and Q-Value system.

2.2 METHODOLOGY

8

Although the most advanced techniques are used in the developed countries, many

nations of the third world like Nepal make use of the direct manual methods. Therefore,

data collection, data transmission, data quality control, storage and retrieval should

always be considered as one consistent information system, and as much attention should

be paid. For hydropower development project, it is important to make sure that series of

appreciable length are available and it is too late to start data collection when the data is

needed. An effective system for routine collection, processing and quality control of data

is therefore and essential part of project. The methodology applied for the study is

grouped into following stages.

2.2.1 Desk study

Topographic maps, aerial photographs, preliminary published and unpublished reports,

journals, field manuals and established theories related to the present study were collected

from the different sources and studied in detail and made the basis for the site

investigation. The toposheet 2858 13 (Sodam) and the aerial photo were used for the

study.

2.2.2 Field study

Field study was carried out in two stages. During first stage survey, the available

information were collected and the information that were collected during desk study

were verified. After reconnaissance study, the plan for the detailed field study was

initiated. Different traverse routes were selected to get the reliable geological information.

The geological, engineering geological and geotechnical data were collected in second

stage. The Brunton Compass, Schmidt hammer, Geological hammer, measuring tape,

dilute HCL (10%), altimeter and different stationary were used for the collection of data.

The main procedure for the collection of primary data in the field are:

1. Detail measurement of the rock discontinuities using the Bronton Compass, Geological Hammer and measuring tape.

2. Measurements of Schmidt hammer value to find out the value of Uniaxial Compressive Strength (UCS).

3. Detail measurement of parameters that is required for rock mass classification using Rock Mass Rating (RMR) and Rock Tunneling Quality Index (Q) System.

9

4. Mapping of rock outcrops, surface deposits and geomorphologic features for the preparation of geological and engineering geological map.

2.2.3 Data processing, Interpretation and Report writing

The data collected during fieldwork were refined and analyzed. Geological map,

engineering geological map, cross section of geological map along headrace tunnel were

prepared using Auto CAD, Arcview 3.2a and Ilwis 3.0. Uniaxial Compressive Strength

(UCS), Rock Quality Designation (RQD), Rock Mass Rating (RMR) and Q values were

analyzed and used for the rock mass classification of the project area. The surface and

underground wedge stability analysis was carried out using DIPS 5.1 and Unwedge

3.005. The geotechnical parameters were calculated using Roclab 1.0. All the data and

maps were then used for the interpretation of geological, engineering geological and

geotechnical condition of the project area. The final report was prepared in accordance

with the guidelines provided by the Central Department of Geology, Tribhuvan

University, incorporating all the analysis, results and data collected in the field

10

CHAPTER THREE

GEOLOGY OF THE STUDY AREA It is generally agreed that the Himalaya is generated as a result of a collision between the

northward moving Indian continent and the Asian landmass. The orogenic process

continues, and mountains are still being formed. Continued activity is manifest in present

day northward movement of the Indian plate at a rate of 5 cm per year and in occurrence

of frequent seismic events along the mountain range and in its surroundings (Seeber and

Armbruster, 1981; Jackson and Bilham, 1994; Pandey et al, 1995; Bilham et al., 1997,

1998). Most of the convergence is accommodated within the Himalaya by movement on

various thrusts and folds. Tectono-morphologically, the whole Himalaya can be divided

into different longitudinal units, each having unique stratigraphic and evolutionary

geological characteristics (Gansser, 1964). From south to north, these units are Sub-

Himalaya, Lesser Himalaya, Higher Himalaya and Tethys Himalaya.

3.1 GEOLOGY OF THE NEPAL HIMALAYA.

The Himalayan Range extends from the Naga Parbat in west to the Namcha Baruwa in

the east with about 2400 km in length and from 200-250 km in width. Nepal Himalaya is

located in the central part of the Himalaya Range and covers about one-third (800 km) of

its total length extending from the Mechi River in east to the Mahakali River in west

(Gansser, 1964). Like the whole Himalayan Range, the Nepal Himalaya is divided into

the five major tectonic divisions. From south to north they are Terai, Sub Himalay, Lesser

Himalaya, Higher Himalaya and Tethys Himalaya (Gansser 1964, Hagen 1969) (Figure

3.1). These zones extend approximately parallel to each other, each characterized by their

own lithology, tectonics, structures and geological history.

3.1.1 Terai Plain

The Gengetic Plain forms the southern fringe of Nepal Himalaya which consists mainly

of alluvial deposits of Pleistocene to Recent age which are derived from the erosion of

sediments from the Himalaya. This zone is separated from the Sub-Himalaya by the

Himalayan Frontal Thrust (HFT) and is the northern edge of the Indo-Gengetic Plain to

the south. The thickness of these deposits is considered to be greatest near the mountain

11

front where mostly gravel and coarse grained sediments are deposited and their grain size

reduces southward where mostly silts and clay are deposited. The average thickness of

deposit is 1500m. Geomorphologically, from north to south the Terai Zone is sub-divided

into Northern (Bhabhar), Middle and Southern zones.

Tibetan Tethys Himalayas

-----------------------------------------STDS

Higher Himalayas

----------------------------------------MCT

Lesser Himalayas

----------------------------------------MBT

Sub-Himalayas

-----------------------------------------HFT

Terai Plain

Figure 3.1: Generalized geological map of Himalaya (Ganser, 1964)

3.1.2 Sub-Himalaya (Siwalik)

The Sub-Himalaya also known as Siwalik is bounded by the HFT in the south and the

Main Boundary Thrust (MBT) in the north. This zone comprises the fluvial deposits of

the middle Miocene to early Pleistone age containing vertebrate fossils (Corvinus, 1988).

This zone is divided into the Lower Siwalik, the Middle Siwalik and the Upper Siwalik

(Auden, 1935). The Lower Siwalik comprises ash grey and red-brown, fine-grained

12

sandstone with pseudo-conglomerate containing pebbles of Siwalik fragments,

interbedded with purple, grey mudstone and siltstone. A few vertebrate fossil remains

have been reported from central and central west Nepal (West et al. 1978, 1981; Munthe

et al, 1983). The Middle Siwalik comprises relatively coarse, arkosic to lithic, grey

sandstone with small proportion of green and grey mudstone and siltstone. Sandstone is

salt-and-pepper type in appearance and is thick-bedded and cross-laminated toward the

top. Occasionally it consists of grit and conglomerate beds in the middle and upper part of

the sequence. Coalified plant logs, leaf impression and some mulluscs are found in

sandstone, mudstone and siltstone (West et al., 1975; Corvinus, 1988). The Upper Siwalik

represents dominant coarse conglomerate beds with minor sandstone and mudstone beds.

Conglomerate consist of pebble, cobble and boulder of gneiss, schist, granite and

quartzite of the Higher Himalaya, limestone, phyllite, slate and sandstone of the Lesser

Himalaya and sandstone of the Lower and the Middle Siwalik. The fragments in

conglomerate can vary in composition from one area to another depending upon the

provenance of the catchment area.

3.1.3 Lesser Himalaya

The Lesser Himalaya is separated form the Higher Himalaya by the Main Boundary

Thrust (MCT) in the north and from the Sub-Himalaya by the MBT in south. The Lesser

Himalaya Zone is characterized by a broad belt of folded and faulted Precambrian to

Pliocene rocks developing a number of thrusts and nappes and mostly comprised of

unfossiliferous sedimentary and metasedimentary rocks such as shale, sandstone,

conglomerate, slate, phyllite, schist, quartzite, limestone, dolomite (Bordet, 1961; Hagen,

1969; Valdiya, 1995; Sakai 1983, 1985; Lefort et al. 1999). There are also some

remarkable granitic intrusions in this zone. This zone is divided into three sub-units

namely Lesser Himalayan Metasedements, Lesser Himalayan Crystallines and Igneous

Rocks. Lesser Himalaya Metasediments are represented by several rock groups such as

Tansen Group, Nawakot Group and Kuncha Group. Lesser Himalayan Crystallines are

represented by Phulchoki Group and Bhimphedi Group. Igneous Rocks are represented

by granites with tourmaline. Tectonically, the entire Lesser Himalaya consists of two

sequences of rocks. They are allochthonous, and autochthonous-parautochthonous units

with various nappes, klippes and tectonic windows. From east to west, the Lesser

Himalaya shows much variation in stratigraphy, structure and magmatism.

13

3.1.4 Higher Himalaya

The Higher Himalaya is bounded by the MCT in the south and the South Tibetan

Detachment Systems (STDS) in the north. It is the hanging wall of the MCT and footwall

of STDS. This zone consists of an approximately 10 km thick succession of crystalline

rocks and comprises mainly high-grade metamorphic rocks such as kyanite-silliminite

bearing gneisses, schist, quartzite and marble of Precambrian age at the basement and

Migmatites and Granites in the upper part. This crystalline unit extends continuously

along the entire length of the country (Heim and Gansser, 1939) commonly called as the

Higher Himalayan Crystalline Series (HHCs). Bordet et al. (1972) divided the Higher

Himalaya into four main units, as Kyanite-Sillimanite Gneiss, Pyroxenic Marble and

Gneiss, Banded Gneiss and Augen Gneiss in an ascending order. However, Le Fort

(1975) divided this zone into three formations as Formation I, Formation II and

Formation III in the ascending order. The Formation I consists of kyanite to sillimanite

garnet, two-mica banded gneiss of pelitic to arenaceous composition. The presence of

augen gneiss, remobilization (migmatization) and intercalation of lime silicate rock and

quartzite characterize the upper part of formation. The Formation II often begins with a

coarse quartzite beds several tens of meters thick. It is mainly composed of alternation of

pyroxene (amphibole) cal-gneiss and marble. The Formation III is characterized by a

more pelitic to greywacke character. The top of the formation is made up of thick coarse

augen gneiss and gradually pass upward to the limestone of Tethyan zone.

3.1.5 Tibetan Tethys Himalaya

The Tibetan Tethys Himalaya lies between the STDS in the south and extends to north in

the Tibet. This is the northern most tectonic zone mainly comprises of fossiliferous

sedimentary rocks such as shale, limestone and sandstone of Late Precambrian-Early

Paleozoic to the Upper Cretaceous (Colchen et al., 1980) deposited in the Tethys Ocean.

This zone is characterized by deep intra-crustal faults, which brought up ophiolites and

melanges within the squeezed sediments that are preserved in several basins of entire

Himalaya. The Tibetan Tethys Unit is exposed in only fewer places within the territory of

Nepal. The rocks of this zone are well exposed and studied in the Thak Khola (Mustang),

Manang, Dolpa, Mt. Everest, Mt. Makalu, Mt Annapurna and Mt. Dhaulagiri.

14

3.2 REVIEW OF PREVIOUS GEOLOGICAL WORK IN CENTRAL NEPAL

A number of geological investigations are being carried since 1875 in Central Nepal at

regional and local scale. The brief description of the geological works is given below.

Medlicott (1875) took a traverse from Amlekhgunj through Kathmandu to

Nuwakot. He was the first to discover the chandragiri-Phulchauki fossiliferrous

beds. He described the sedimentary and the low grade metamorphic rocks to the

south and gneiss and the high grade rocks to the north of Kathmandu.

Auden (1935) carried the first systematic geological investigation in Nepal, who

visited some parts of the Eastern and the Central Nepal. He gave a fairly good

account of the geology of this part of the Himalaya. He studied the fossils from

limestone of Chandragiri and assigned the Ordovician age to these rocks. He

noticed superposition of the high grade metamorphic rocks over the low grade

metamorphic rocks in the Mahabharat range.

Bordet et al. (1960) did pioneer works in Nepal. They described the geology of the

Phulchauki area and confirmed the Silurian age of the rocks at the base of

Phulchauki on the basis of trilobite fossils. They further extended their work in the

Central-West Nepal in Pokhara region (Bordet et al 1964, 1972) and their

contribution to the geology of this region is noteworthy.

Gansser (1964) compiled the geology of Nepal and tried to reconstruct the

comprehensive and total geological configuration of the Himalaya. In his work, he

has tried to give a regional tectonic outline of the whole Himalaya including

Nepal.

Hagen (1969) worked the first most important and extensive study on the Nepal

Himalaya. He developed the concept of nappe structures in the Nepal Himalaya.

He divided the geology of the study area into two nappes, the Kathmandu Nappe

and the Nawakot Nappe. The distinction of these was based on conspicuous

differences in composition, metamorphic grade and age. Based on the lithological

comparisons with the Alps, he placed the Nawakot Nappe in the Paleozoic-

Mesozoic and thus considered it as the younger than the overlying Kathmandu

Nappe of the Precambrian–Early Paleozoic age. He proposed that the relatively

high grade metamorphic rock of the Kathmandu Nappe is tectonically

emplacement over the relatively low grade metamorphic rock of the Nawakot

15

Nappe. The Kathmandu Nappe was interpreted as an erosional relict of a once

extensive thrust sheet, rooted in the central crystalline and still linked with it by a

‘tectonic bridge’ formed of the Gosaikund Gneiss north of Kathmandu.

Hashimoto et al. (1973) divided the Kathmandu region into the Gosaikund gneiss

zone, Main Central Thrust zone, Nawakot metasediment zone, Sunkoshi tectonic

zone, Sheopuri injection gneiss zone, Kathmandu basin, Granite intrusion zone,

Mahabharat zone and the Siwalik zone based mainly on correlation of the

lithostratigraphy. They have sub-divided the thick sequence of midland

metasediment group into four divisions based on lithologic characters as the

calcareous succession, siliceous succession, arenaceous succession and the

argillaceous succession.

Sharma (1973) divided the Himalaya of Central Nepal into a Basement Gneiss

Complex above the Main Central Thrust (MCT), the Mahabharat Limestone

Group, the Phulchauki Formation, the Chandragiri Formation and the Churia

Group.

Stöcklin and Bhattarai (1977) studied the geology of the Kathmandu area and

Mahabharat range based mainly on the photo-geological interpretations supported

by field works. They developed the stratigraphy of the Central Nepal. Apart from

the tertiary Siwalik and Quaternary deposits, they grouped the rocks into two

largest units, Nawakot Complex and Kathmandu Complex. These complexes are

further sub-divided into formations and members.

Stöcklin (1980) noted that the crystalline complex of Kathmandu consists

primarily of a right-way-up sequence of regionally metamorphosed sediments

displaying a metamorphic zonation roughly concordant with stratigraphy with a

gradual decrease in metamorphic grade from garnet-schist at the base to barely

metamorphosed, fossiliferous Paleozoic sediments on top. The contact of the

Kathmandu Crystalline Zone with the underlying Nawakot Meta-sediments is

marked by both intense shearing and by a stratigraphic, metamorphic and

structural discontinuity indicating a thrust plane. The Kathmandu Crystalline Zone

is interpreted as the remnant of a nappe, rooted in the main part of the Higher

Himalayan Crystalline.

16

Rai (1998) mentioned that the Kathmandu–Gosainkund region can be divided into

the Gosainkund Crystalline Nappe (GCN) and the Kathmandu Crystalline Nappe

(KCN). The boundary between the two nappes is MCT. The GCN and the KCN

are thrust over the Lesser Himalaya along the MCT and the Mahabharat Thrust

(MT) respectively.

Upreti (1999) divided the crystalline nappes of the Lesser Himalaya into two

groups depending upon the stratigraphy and metamorphism of rock units: 1.

Nappes composed of upper amphibolite to granulite facies rocks, similar to rocks

of the Higher Himalayan Zone or Tibetan Slab. 2. Nappes of the Bhimphedi

Group (composed of the low- to medium-grade metamorphic rocks such as

biotite-garnet-schist and marble) with the Lower Paleozoic cover. The Kathmandu

Nappe was placed in the second group.

Upreti & Le Fort (1999) proposed that subsequent to the MCT another thrust

Mahabharat Thrust (MT) is developed further south which took up the movement

that had earlier been on the MCT. The MT carried the rocks of the Bhimphedi

Group to their present position on the top of the Lesser Himalayan rock. A

combination of movements along the MCT and the MT may have served to bring

the whole Bhimphedi Group to the surface.

Johnson et al. (2001) refused the concepts that Kathmandu complex is a Klippe or

separate thrust sheet. They showed it is ductile shear zone or ductile fault zone

with no obvious single break between the Nawakot and Bhimphedi Groups.

Instead, there is a zone of mylonites and phyllonites about 1.5 km thick.

Acharya et al. (2007) carried out the micro-structural analysis in the western part

of the Kathmandu Nappe (Galchhi and Malekhu area) and mentioned that the

petrography, metamorphism and nature of strain of both hanging and footwall

rocks is strikingly consistent with description of the MCT elsewhere. They

conclude that the thrust belt is MCT and the Kathmandu Nappe is required to be

an MCT re-entrant.

The detailed geological study of the project area is done by limited number of geologists

and agencies. The published or unpublished report of the study area is hard to find and

only limited number of papper are collected. Besides these, following are the important

work done in the study area.

17

Medium Hydro Power Study Project, NEA (1998) had carried out the reconnaissance

study of Upper Trisuli 3 (UT-3) which is named as Gogane to Betrawati Hydroelectric

Project and Upper Trilsuli 3A (UT-3A) is mutually included in UT-3 Project.

Geological and Geotechnical study of Upper Trisuli - 3A hydroelectric project conducted

by Soil, Rock and Concrete Laboratory, Engineering Services, Nepal Electricity

Authority.

Department of Mines and Geology has also prepared and compiled regional geological

map including the project area in the scale of 1:1,000,000 (1994). According to DMG, the

study area lies in the Ulleri Formation.

Upreti B. N. (1999) has placed the present study area in northern section of the Gorkha-

Nawakot Metasediment Zone within the Ulleri Formation.

3.3 GEOLOLOGY OF THE CENTRAL NEPAL

The geology of the Central Nepal, Lesser Himalaya includes the area between the Dudh

Kosi River in the east and the Marsyangdi River in the west. The area consist of exposed

sedimentary and metamorphic rock sequences in a wide zone and complicated by the

presence of folds, thrusts and imbricated zones (Figure 3.2).

From east to west the Central Nepal Lesser Himalaya may be divided into three

transverse zones (Upreti, 2000). They are

I. Chautara-Okhalhunga Metasediment Zone

II. Kathmandu Nappe

III. Gorkha-Nawakot Metasediment Zone.

The Kathmandu Nappe with crystalline rocks tectonically overlies Chautara-Okhaldunga

and Gorkha-Nawakot metasediment zones. Stratigraphically, Gorkha-Nawakot Zone and

Chautara-Okhaldunga Zone belongs to the same stratigraphic unit separated by

Kathmandu Nappe.

18

Figure 3.2: Geological map of the Central Nepal Himalaya after Colchen et al., 1986,

Modified by Rai, 2001

3.3.1 Chautara-Okhaldunga Metasediment Zone

Chautara-Okhaldunga Metasediment Zone mainly occupy the area along the E-W stretch

of the Sunkosi River and lying to the east of Kathmandu Nappe and north of the narrow

arm of the outlying crstalline klippen of the Mahabharat Range, predominantly composed

of metasediments along with a thick sequence of augen gneisses in the northern part. The

metasediments are composed of slates, phyllite and metasandstones of low-

grademetamorphism, whereas the metasediment lying between the augen gneiss and the

MCT are crystalline schists and phullites, quartzites and calc-schists and limestones

(Kano, 1984).

From south to north, the stratigraphy and structures of the eastern part of the Lesser

Himalaya of Central Nepal may be described according to following tectonic zones

(Ishida and Ohata, 1973; Maskey, 1986; Sharma, 1973; Schelling, 1989).

19

North

Higher HimalayanCrystalline Zone (Tibetan Slab)

-------------------Main Central Thrust (MCT) -----------------

Jiri Metasediment Zone

----------------------------------------------------------------------

Melung-Salleri Augen Gneiss Zone

---------------------------------------------------------------------

Chautar-Okhaldunga Metasediment Zone

-------------------------------Thrust--------------------------------

Gondwana Zone ( Tansen Unit)

----------------------------------Thrust-------------------------------

Mahabharat Crystalline Zone (Kathmandu Nappe)

----------------------------------Thrust--------------------------------

Southern Metasediment Zone

------------------Main Boundary Thrust (MBT) -------------------

Churia Group (Siwaliks)

South

3.3.2 Kathmandu Nappe

The Kathmandu Nappe was first recognized by Hagen (1969) and later mapped in detail

by a team of geologist of the Mineral exploration Project from the Department of Mines

and Geology, Nepal (Stocklin and Bhattarai, 1977; Stocklin, 1980). The rocks of

Kathmandu Nappe have been included into Kathmandu Complex which is further divided

into two groups; the Precambrian Bhimphedi Group consisting of relatively high grade

metamorphic rocks, and the Phulchauki Group of unmetamorphosed to weakly

metamorphosed sediments containing fossil of Lower-Middle Paleozoic age. The name of

these formations proposed by Stocklin and Bhatarai, 1977; Stocklin, 1980 with their main

lithology and thickness is shown in Table 3.1.

20

3.3.3 Grokha-Nawakot Metasediment Zone

The Gorkha-Nawakot Metasediment Zone occupies the low-grade metasedimentary rocks

outcropping the north, west and southwest of Kathmandu Nappe which has been grouped

into the Nawakot Complex (Stocklin and Bhattarai, 1977; Stocklin, 1980). The rocks of

Nawakot Complex have been subdivided into the Lower and Upper Groups separated by

an erosional unconformity (Table 3.1).

Table 3.1: Stratigraphy of Kathmandu Complex (Stocklin and Bhattarai, 1977; Stocklin,

1980)

Rock Unit Group Formation. Thickness (m) Main Lithology

Kathmandu Complex

Phulchauki Group

Godavari Limestone 300 Limestone, dolomite Chitlang Formation 1000 Slate

Chandragiri Limestone 2000 Limestone

Sopyang Formation 200 Slate, calc-phyllite Tistung Formation 3000 Metasandstone, Phyllite

----------------------------Transition-------------------------------------------

Bhimphedi Group

Markhu Formation 1000 Marble, schist Kulekhani Formation 2000 Quartzite, schist

Chisapani Quartize 400 White quartzite

Kalitar Formation 2000 Schist, quartzite Bhainsedobhan

Marble 800 Marble

Raduwa Formation 1000 Garnet-schist, quartz

---------------------------------------Mahabharat Thrust---------------------------------

Nawakot Complex

Upper Nawakot

Group

Robang Formation 200-1000 Phyllite, quartzite

Malekhu Formation 800 Limestone, dolomite

Benighat Slate 500-3000 Slate, argillite, dolomites

--------------------------Erosional Unconformity (?)-------------------------

Lower Nawakot

Group

Dhading Dolomite 500-1000 Stromatolitic dolomite

Nourpul Formation 800 Phyllite, metasandstone, dolomite

Dandagoan Phyllite 1000 Phyllite

Fagfog Quartzite 400 White quartzite

Kuncha Formation 3000 Phyllite, quartzite, gritestone

21

3.4 GEOLOGY OF THE PROJECT AREA.

The geology of the project area has been studied from Mailung Dobhan in north to

Archale in south. The project area lies to Kuncha Group of Lesser Himalaya

Metasediments in Central Nepal. Geologically, the study area is represented by two unit;

Schist Unit and Gneiss Unit. The Main Boundary Thrust (MBT) is located at about 80 km

south of project area and Main Central Thrust (MCT) about 25 km north of project area.

A local anticline is expected to exist in the project area whose axis lies in the tunnel

alignment. In general, the foliation of rocks within the project area dips NE-NW and SE-

SW with dip amount of 10º-30º.

3.4.1 Schist Unit

This unit is exposed around Mailung Dobhan, downhill of Chepleti gaun, Danda gaun,

Diyale and Khadku gaun. Schist Unit consists of light grey, medium- to thick-banded

psammatic schist with occasional band of medium- to thick-banded (up to 300 cm) pelitic

schist and medium-banded (30 cm to 100 cm) quartzite. Quart veins of thickness up to 7

cm are also observed. It is well exposed at both banks of the Trisuli River at headwork

site. It is slightly to moderately weathered, medium strong and seamy to blocky.

Along the right bank of the Trisuli River from the Chipleti Khola to Mailung Dobhan,

road cut sections and cliff of schist are observed. The exposures consists predominantly

grey psammatic schist with occasional beds of pelitic schist and quartzite. Quartz veins of

thickness up to 7cm are also present. Quartz veins show folded boudinage parallel to the

foliation. Attitude of foliation is 143º/16º NE, 168º/15ºNE, 105º/20º NE, 100º/14º NE,

235º/17º NW and 240º/28º NW.

Along the Mailung Khola from Mailung Dobhan to confluence between the Mailung

Khola and the Nyam Nyam Khola, the right bank of the Mailung Khola is covered by

alluvium and colluviums while left bank show the steep exposure. Lithology is light grey,

massive, thick-banded psammatic schist with parting of pelitic schist. Attitude of foliation

is 139º/22º NE and 156º/29º NE. The contact between schist and gneiss is observed at left

bank of the Mailung Khola near the half the way between Mailung Dobhan and

confluence between the Mailung Khola and the Nyam Nyam Khola.

22

Along the right bank of the Trisuli River along the foot trail towards Gogane gaun, dark

grey, blocky psammatic schist inter-banded with pelitic schist is observed. Attitude of

foliation is 139º/22º NE and 219º/09º NW. A contact between gneiss and schist is

observed at right bank of the Trisuli River, about 150 m uphill from confluence between

the Mailung Khola and the Trisui River, at altitude of 1300 m. Attitude of foliation at

contact is 130º/30º NE.

Along the foot trail from Simle, Danda gaun to Khadku, Simle to Danda gaun section is

totally covered by colluviums and is cultivated. Few exposures are observed along Danda

gaun to Khadku section along the tributaries of the Trisuli River. The lithology is dark

grey, massive psammatic schist with occasional bands of pelitic schist. Quartz veins of

size up to 5 cm are also observed. Attitude of foliation is 160º/18º NW, 158º/17º SW and

138º/14º SW.

Along the Dharni Khola, a contact between massive white augen gneiss and light grey,

medium-banded, fractured, psammatic schist is observed. Attitude of foliation is 195º/26º

NW.

Along the Chipleti Khola from confluence between the Chipleti Khola and the Trisuli

River to Chipleti gaun, light grey, massive schist with occasional bands of quartzite is

observed. Quartz veins of thickness up to 5cm are also observed. Attitude of foliation is

123º/22º NE and 105º/20º NE. A contact between white, massive, thick-banded gneiss

and light grey, medium-banded psammatic schist is observed at Chipleti Khola, at altitude

of 1080 m, about 20 m from the Shree Chipleti Primary School. Attitude of foliation is

128º/15º NE.

3.4.2 Gneiss Unit

This unit is exposed around Simle, downhill of Diyale, Siruchet and uphill of Chipleti

gaun. This unit comprises of milky white, medium- to thick-banded augen gneiss with

occasional parting of light grey to greenish grey schist. Quartz veins up to 5 cm are also

observed in gneiss. Only few exposure of gneiss band is observed with in the study area.

The gneiss is generally slightly to moderately weathered, medium strong to strong in

strength and blocky to massive.

23

Along the road cut section from Simle to the Chipleti Khola at right bank of the Trisuli

River, milky white, massive, thick-banded augen gneiss with the parting of pelitic schist

is observed. Quartz veins of thickness up to 10 cm are also observed. Attitude of foliation

is 095º/22º SW, 195º/18º SE, 118º/16º SW and 110º/15º SW. A contact between augen

gneiss and schist is observed at downhill of Diyale gaun along road cut section. Granite of

thickness up to 40 cm intruded in both gneiss and schist bed is observed. The intrusion is

parallel to the foliation of beds of gneiss and schist. This contact area shows the core of

the local anticlinal structure. The schist bed dips toward northwest while gneiss bed dips

towards southwest. Attitude of schist and gneiss bed are 240º/28º NW and 110º/15º SW

respectively.

Along the foot trail from Khadku to Siruchet, white, massive, thick-banded augen gneiss

is observed. Quartz veins are also present. Attitude of foliation is 108º/07º NE, 096º/24º

NE and 116º/07ºNE.

At the confluence between Nyam Nyam Khola and Mailung Khola, at right bank of the

Mailung Khola, white, massive and blocky, augen gneiss with four distinct joint sets is

observed. Attitude of foliation is 017º/46º NW and 110º/14º NE. Attitude of joints are;

J1: 094º/68º, J2: 177º/87º and J3: 130º/39º.

25

CHAPTER FOUR

SEISMICITY OF THE PROJECT AREA

4.1 SEISMICITY OF NEPAL

Earthquake generation is confined to the crustal depth of about 20 km. It is generated as a

result of released stresses, which are accumulated in the geodynamic under thrusting

process of the Indian plate against the Eurasian plate. However the shallow earthquakes

of a depth up to 6 km are generated as a result of strike slip faults.

The records of seismic activities are limited in the Nepal Himalayas and hence correlation

of seismic events with the adjacent Himalayan Region would be a useful source of

information for designing the hydraulic structures. Several seismicity studies have been

carried out for various projects in the country during the study and engineering design

phases.

4.2 SEISMICITY EVALUATION

Nepal has experienced a number of large earthquakes over the past few decades which

have caused the substantial damage of life and property. A micro seismic epicenter map

of Nepal Himalaya and adjoining region (1:200000) is prepared by the National

Seismological Center, Department of Mines and Geology (DMG) (Figure 4.1). The map

shows the distribution pattern of earthquake epicenter in Nepal and adjoining region. The

map also suggests that Far Western Nepal is seismically more active than Eastern Nepal.

It is also clear from the map that there is a dense cluster of earthquake epicenters in Far

Western Nepal, less in Eastern Nepal and the least in Central Nepal.

There are several methods to convert the maximum acceleration of the earthquake motion

into the design seismic coefficient. Simplest method, empirical method and dynamic

analysis using dynamic model are common methods to establish the seismic coefficient.

The simplest method is represented by

α = Amax/980

26

Where,

α = Design seismic coefficient

Amax= Maximum acceleration of motion (gal).

However, this method will evaluate rather larger value of seismic coefficient compared

with real value.

The Empirical method is represented by

αeff = R α = R Amax/980

Where,

αeff = Effective design coefficient.

R = Reduction factor (empirical value, R = 0.5-0.65)

This method is considered to be most common method to establish the design seismic

coefficient.

Dynamic Analysis Method using Dynamic model requires lots of parameters like design

input motion, soil structure model, properties of rock materials etc. Therefore detail study

is required to use this method.

4.3 NEPALESE STANDARD

In order to determine seismic coefficient, a seismic design code for Nepal has been

prepared. The country is divided into three seismic risk zones based on allowable bearing

capacity of three types of soil foundation. The Upper Trisuli - 3A HEP is located in the

third seismic zone of Nepal (Figure 4.3), and the soil foundation at dam site belongs to

average soil type. Therefore, the basic horizontal seismic coefficient is considered to be

0.08. By using Empirical method, the effective design coefficient according to seismic

design code of Nepal is given by

αeff = R α = R Amax/980

Where,

αeff = Effective design coefficient.

R = Reduction factor (empirical value, R = 0.5-0.65)

27

For maximum acceleration of 250-300 gal according to Seismic Hazard Map (Figure 4.2),

published by DMG, National Seismological Center, and the reduction factor 0.5, the

calculated effective design seismic coefficient for Upper Trisuli - 3A HEP is

approximately 0.13 to 0.15.

Figure 4.1: Micro seismicity epicenter map of Nepal (prepared by National

Seismological Center/Department of Mine and Geology).

4.4 INDIAN STANDARD

In order to determine the design horizontal coefficient, a seismic risk map for India has

been prepared. The map is published in Indian Criteria for Earthquake Resistant Design

of structure. The country is divided into five seismic zones in India Standard (Figure 4.4).

According to seismic risk map of India, Nepal lies in fifth seismic risk zone of India

(zone V). Therefore, it can be considered that Upper Trisuli Hydroelectric Project is

located in the fifth seismic zone of India (zone V) and the basic horizontal seismic

coefficient (αo) can be taken as 0.08.

The design horizontal seismic coefficient in the Indian Standard is defined by the

equation

28

αh = β I αo

Where,

αh = Design horizontal seismic coefficient

β = soil foundation factor (1 for dam)

I = Importance factor (2 for dam)

αo = Basic horizontal seismic coefficient

Therefore, the design horizontal coefficient for Upper Trisuli - 3A Hydroelectric Project

for dam is 0.16 according to Indian Standard.

Figure 4.2: Seismic hazard map of Nepal (published by Department of Mines and

Geology)

29

Figure 4.3: Seismic risk map of Nepal (source: Kaila K.I., Gaur U.K. and Narain, H

(1972)).

30

Figure 4.4: Seismic hazard map of India (Source: Kalia, k.l, Gaur U.K. and Narain, H

(1972))

31

CHAPTER FIVE

ENGINEERING GEOLOGICAL INVESTIGATION

OF THE PROJECT AREA Engineering geological studies are the most essence work to provide actual geological

conditions of the area. It includes engineering geological mapping of the major hydraulic

structures of the project and rock mass classification of the headrace tunnel area with the

stability analysis and the preliminary support design. Statistical joint analysis of the

headrace tunnel has been done on the basis of the detail measurement of discontinuities.

Rock mass rating (RMR) and rock tunneling quality index (Q) are used for the rock mass

classification which helps to study the characteristics and quality of rock mass of the

proposed headrace tunnel.

5.1 ENGINEERING GEOLOGICAL CONDITION OFTHE HEADWORKS

The proposed headwork site is located at about 1000 m downstream from confluence of

the Mailung Khola and the Trisuli River (Figure 5.1). The headwork comprises the

diversion weir, intake canal, aquaduct and desanding basin. The engineering geological

map of headworks is presented in Figure 5.2.

(a)

32

(b)

Figure 5.1: (a) Upstream view and (b) Down stream view of the headwork site.

5.1.1 Diversion Weir

The diversion weir is located at straight course of the Trisuli River. The Trisuli River

flows in direction 150º and has width of about 25 m at weir axis. The shape of valley at

diversion axis is V shaped. Both bank of the Trisuli River at weir axis are represented by

terrace deposit consisting of both alluvium and colluviums (Figure 5.3). The alluvium and

colluviums deposit consist of boulder, cobbles and pebbles of schist and gneiss with silty

and sandy matrix. The average slope of left bank of the river valley near the weir axis is

about 20º and that of right bank is about 15º. The bed rock has been observed at about 50

m away from the right abutment of river and it is about 125 m away from the left

abutment of the river. The outcrop is continuous in both banks and consists of light grey,

slightly to moderately weathered, medium in strength, thin- to medium-banded schist

with occasional bands of light grey fined grained quartzite. The foliation planes are

dipping NE-NW with amount ranging 10º to 30º. Other two sets of joints dipping NE and

SE have been observed. These joints are moderately open, rough to irregular, moderately

spaced with moderate persistency. The strike of the foliation plane is almost

perpendicular to river flow which is favorable orientation for construction of weir.

The detailed surface joint mapping has carried out in the rock exposure around diversion

weir. The statistical analysis shows the following common joints.

Dip/Dip direction

23º/007º

78º/049º

68º/105º

Figure 5.4: Stereographic projection of discontinuities measured around Diversion Weir

The value of RMR ranges from 35-70 and Q value from 0.88 to 5. The value of these

indices indicates that the rock mass around diversion weir is categorized as class II to

class IV which is defined as good to poor rock.

The diversion weir has been investigated by two drill holes; DH-1 and DH-2, and seismic

refraction survey (Annex II). DH-1 and DH-2 drill holes lie at the left and right bank of

the Trisuli River along weir axis. The bed rock in river channel could not be obtained up

to 35 m in drill holes. So the bed rock is estimated to be more than 35 m below river

level. The seismic refraction survey also shows that the depth of bed rock at weir axis is

more than 30 m.

The stability of the slope at both banks at dam site is generally favorable due to upstream

dipping of rock. No slope failures except minor surface erosion are observed. Except this;

there are some prominent gully erosions. The major gully at the headwork site is the

37

Dharni Khola. It seems that it has the capability to bring huge amount of boulders as they

are lying on its course. Though during field study, it was dry but at rainy season it could

be very dreadful.

5.1.2 Intake Canal

Intake canal is proposed to convey water from intake into desander. The proposed intake

canal passed through alluvium terrace deposit below the slope of colluviums terrace. The

alluvium terrace deposit consists of rounded to sub-rounded boulder, cobble and pebbles

of schist, quartzite and gneiss in sandy and silty matrix. The canal passes through a major

gully the Dharni Khola. Lots of huge boulders are lying on its way. So, serious attention

is necessary to prevent the possible danger to control.

5.1.3 Aquaduct

10 m long aquaduct has been proposed across the Dharni Khola to protect the canal

structure from possible debris flow through the Dharni Khola. The aquaduct will be

founded on terrace deposit consist of rounded to sub-rounded boulder, cobble and pebble

of schist, quartzite and gneiss in sandy and silty matrix.

5.1.4 Desanding Basin

The proposed desanding basin lies on the right bank of the Trisuli River on the alluvium

deposit. The thickness of alluvial deposit from river bed level to top is about 20 m and has

enough space for surface desander. The alluvial deposit comprises of cobbles and pebbles

of schist and gneiss with silty, sandy and clayey matrix. Adjacent slope is steep and

consist of an older rock fall deposit. This rock fall consist of huge boulders of schist and

gneiss of size up to 5 m. It seems that the rock fall had occurred in the past in more than a

single phase. Special attention shall be taken in future from the stability point of view to

this rock fall.

The surface material around the desanding basin is covered by alluvium and colluviums.

The desander area has been investigated by drill hole DH-3 and seismic refraction survey.

The drill hole and seismic refraction survey have shown that the thickness of overburden

deposit at desander area is more than 25 m as the rock could not be encountered at this

depth by drilling. So, the structure will be solely founded on alluvial deposit. The

constant head test conducted in this drill hole showed the permeability of alluvial deposit

38

varies in the range 5.07 ×10-3 cm/sec to 2.22 ×10-2 cm/sec (Annex II). So the alluvial

deposit at desander basin area is highly permeable.

5.2 ENGINEERING GEOLOGICAL CONDITION OF DIFFERENT

STRUCTURES

5.2.1 Intake Portal

The intake portal is located on rocky cliff along the right bank of the Trisuli River (Figure

5.5). The rock at the portal area is composed of light grey, slightly weathered, medium in

strength schist with thin intercalation of quartzite whose thickness varies up to 5 cm.

Three prominent joint sets are observed in intake portal site; F: 350º/19º, J1: 029º/55º and

J2: 093º/52º. They are close to moderately close spaced, tight to moderately open

aperture, fresh to slightly weathered, medium persistence, planar to rough surface with

coating of sandy and silty material. The wedge stability analysis at intake portal is carried

out by graphical method and shown in Figure 5.6.

The RQD, RMR and Q value are 58%, 53 and 4.58 respectively. The RMR value

suggests that the rock is fair where as Q value suggest that the rock is poor. The rock at

intake portal dips towards NW-NE with amount ranging from 10º to 30º. The natural hill

slope at intake portal is 70º towards east. The joint J2 dips towards the natural hill slope.

This orientation is less favorable for the cut slope. There is a possibility of plane failure

due to joint set J2 towards east and a wedge failure due to joint sets J1 and J2 whose dip

and dip direction is 40º/065º (Figure 5.6). Therefore, the cut slope is to be treated properly

by rock bolts and shotcrete.

39

Figure 5.5: Photograph of Intake Portal facing 250º

Figure 5.6: Stereographic projection of discontinuities measured at Intake Portal

40

5.3 Engineering Geological Condition of Headrace Tunnel

The proposed headrace tunnel is about 4142 m and will have excavated diameter 6.2 m. It

will be D shaped and passes through the right bank of the Trisuli River at an average

elevation of 875 m. The maximum rock cover is about 300 m at chainage 3+470 m and

the minimum rock cover is about 90 m at chainage 1+907 m and 2+923 m.

The outcrops exposed along the tunnel alignment were mapped extensively and detail

joint measurements were taken. Since the most of the tunnel passes under thick

colluviums, a result of mapping along rivers, streams, gullies and foot trails was projected

to the tunnel horizon in order to produce the required geological information along the

tunnel route. It should be noticed that the geological condition along the tunnel alignment

is largely based on surface mapping. Considering all parameters engineering geological

map of the headrace tunnel area has been prepared at 1:15000 scale and shown in Figure

5.7. The geological cross section has been prepared along the headrace tunnel with

tentative support pattern and shown in Figure 5.8. The statistical joint analysis is done as

far as possible. The stability analysis is done based on the stereographic projection

assuming the friction angle 35º for rock. RMR and Q are used for rock mass

classification.

The headrace tunnel passes through mainly two types of rocks namely schist with thin

intercalation of quartzite and augen gneiss. The schist will occupy about 5% of the tunnel

length and the rest 95% will be occupied by augen gneiss. One anticlinal axis is expected

in the tunnel alignment and no other structures such as fault and thrust are noticed in the

tunnel alignment. In general, the rock along the tunnel is considered to be medium strong

to very strong in strength. The rock is slightly to moderately weathered. Gneiss is stronger

in strength as compared to schist. The rock is exposed mainly in the small creeks and at

higher elevation in the form of steep cliff along the tunnel routes. No major fault crossing

the tunnel are noticed during the field mapping however several thin bands of fracture

zones are noticed in tunnel zone mainly along the tributaries and at anticlinal axis. The

mapping in the river section was projected to the tunnel horizon in other to produce

required geological information along tunnel route.

Most of the tunnel length passes nearly perpendicular to the strike of foliation of rock

dipping 10º-30º due NE-NW (at intake site) to SE-SW (at power house site). This

41

orientation is generally considered to be favorable tunneling condition for excavation if

tunnel drive with dip.

A discontinuity survey was carried out in several directions on the different rock exposure

along the headrace tunnel alignment corridor, on the slopes and along the small creeks

and streams, and observed data are statistically analyzed. Joint mapping revealed mainly

three sets of joints along the tunnel with some random sets. The joints are tight to

moderately open, close to moderately spaced, continuous (3-10 m), rough irregular and

occasionally smooth to planner with silt and sandy coating surface. As the tunnel passed

through anticline, the orientation of foliation joint along tunnel alignment become

different. Towards headwork site, the foliation has NE-NW dip direction and towards

power house site it has SE-SW dip direction. The engineering geological condition of the

headrace tunnel at different chainage is presented in following paragraphs.

Ch 0+000 m to Ch 0+100 m

The tunnel alignment from Ch 0+000 m to Ch 0+100 m runs with an azimuth 30º. The

surface topography is steep. The rock mass is light grey, medium-banded, unweathered to

slightly weathered, medium strong psammatic schist with occasional bands of quartzite

up to 5cm. Joints are close to moderately close spaced, tight to moderately open aperture,

medium persistence, rough, undulating surface with silty coating surface. The ground

water condition is dry. The attitude of joints are F: 010º- 306º/14º - 24º, J1: 117º-

142º/62º-70º and J2: 035º-055º/59º-65º. The RQD, RMR and Q are 64%, 53 and 5.83

respectively which suggest the rock mass is in III class and described as fair rock.

Ch 0+100 to C h 0+200 m

The tunnel alignment from Ch 0+100 m to Ch 0+200 m runs with an azimuth 30°. The

rock mass is light grey, thin to medium banded, unweathered, low strength psammatic

schist with parting of quartzite. The joints are closely spaced, moderately open, medium

persistence, rough to planar with silty coating surface. The ground water condition is dry.

Attitude of joints are F: 300º/17º, J1: 020º/70º and J2: 079º/41º. The RQD, RMR and Q

values are 45%, 35 and 3.75 respectively which suggest the rock mass is IV class and

described as poor rock.

Figure 5.9: Stereographic projection of joints measured along road cut section at Chepleti.

Ch 0+200 m to Ch 0+707 m

The tunnel alignment from Ch 0+200 m to Ch 0+707 m runs with an azimuth 30º.The

rock mass is milky white, medium- to thick-banded, unweathered to slightly weathered,

medium to high strong augen gneiss with parting of schist. Joints are wide to very wide

spaced, tight to moderately open aperture, medium persistence, rough and irregular

surface with silty coating surface. The ground water condition is dry. The attitude of

joints are F: 013º-031º/09º-14º, J1: 129º-197º/56º-86º: and J2: 047º-090º/70º-75º. The

RQD, RMR and Q are 80%, 66 and 13.69 respectively which places the rock mass in II

class and described as good rock.

Ch 0+707 m to Ch 0+876 m


rock mass is milky white, medium-banded, slightly weathered, medium strong augen

gneiss. Joints are moderately close spaced, moderately open aperture, medium

persistence, planar to rough surface with silty coating. The ground water condition is

completely dry. The attitude of joints are F: 330º/28º, J1: 272º/68º: and J2: 068º/31º. The

RQD, RMR and Q are 45%, 35 and 3.75 respectively which places the rock mass in IV

class and described as poor rock.

45

Ch 0+876 m to Ch 1+938 m


rock mass is milky white, medium- to thick-banded, slightly weathered, strong augen

gneiss. Joints are moderately close spaced, moderately open to open aperture, medium

persistence, planar to undulating surface with silty coating. The ground water condition is

completely dry. The attitude of joints are F: 221º/13º, J1: 024º/60º: and J2: 099º/71º. The

stereographic projection of joints measured along the road cut section is shown in Figure

5.10. There are three wedges formed. W1 is in daylight condition and W2 and W3 seems

to be stable since they are gentle than friction angle. The RQD, RMR and Q are 58%, 53

and 4.58 respectively which places the rock in III class and described as fair rock.

Figure 5.10: Stereographic projection of discontinuities along road cut section at downhill

of Katunje guan.

Ch 1+938 m to Ch 2+600 m



gneiss with parting of schist. Joints are wide to moderately close spaced, moderately open

to tight aperture, medium persistence, planar to undulating surface with silty coating. The

ground water condition is moist. The attitude of joints are F: 210º/14º, J1: 130º/54º: and

J2: 183º/79º. The stereographic projection of joints measured along the road cut section is

46

shown in Figure 5.11. There are three wedges formed. W1 is in daylight condition and

W2 and W3 seems to be stable since they are gentle than friction angle. The RQD, RMR

and Q are 73%, 65 and 9.90 respectively which places the rock in category II to III class

and described as good to fair rock.


downhill of Diyale guan.

Ch 2+600 m to Ch 4+142 m



gneiss. Joints are wide to moderately close spaced, moderately open to tight aperture,

medium persistence, planar to undulating surface with silty coating. The ground water

condition is moist up to the Ch 3+200 m and rest part is dry. The attitude of joints are F:

189º/14º, J1: 020º/67º: and J2: 113º/45º. The stereographic projection of joints measured

along the road cut section is shown in Figure 5.12. There are three wedges formed. W1 is

in daylight condition and W2 and W3 seems to be stable since they are gentle than

friction angle. The RQD, RMR and Q are 70%, 63 and 7.48 respectively which place the

rock in category II to III class and described as good to fair rock.

47


at down hill of Danda guan.

Rock mass condition in the headrace tunnel has been based on geological mapping and

detail joint mapping on surface rock outcrops. Geomechanical classification using both

rock mass rating (RMR) (Bieniawski, 1989) and rock tunneling quality index (Q) (Barton

et al, 1974) has been carried out. All parameter of rock mass classification are taken in the

field. Parameter used for rating assigned different values in different condition of

discontinuity. Detail record of data sheet is presented in Annex V. Adjusted value of

RMR (adjustment made taking into account of the tunnel orientation with respect to

discontinuities) and Q value are separately used in classification of rock mass. The RMR

of the tunnel area ranges from 25 to 70 and Q values from 3 to 10 and summarized in

Table 5.1 and Table 5.2

5.4 Engineering Geological Condition of Surge Tank

The proposed underground surge tank is located on the right bank of the Trisuli River.

The surge tank will be of about 35 m high having excavated diameter of about 17 m. The

general natural ground slope at the surge tank area is about 60º. The surface area of surge

tank is covered by colluviums deposit, which consists of the boulders of gneiss in sandy

and slity matrix. The maximum size of boulder up to 5 m lying on the slope is observed.

The surge tank area is investigated by drill hole DP-4. A borehole DP-4 has been drilled

up to 60.250 m depth in the surge tank area. The drill hole showes fresh to slightly

48

weathered gneiss with very poor to very good RQD value. In general the RQD value

obtained is fair to poor. The RQD value is 43% for the whole length of borehole. The

thickness of colluviums in the surge tank area is about 18 m according to drill hole DP-4.

The overburden consists of colluvial material in the slope underlying the gneiss. The

RMR and Q values obtained from the surface mapping at outcrop around surge tank area

showed 61 to 65 and 5.83 respectively which places the rock in class II to III type with

description of fair rock quality.

Table 5.1: Rock mass classification along headrace tunnel based on RMR value. Chainage Length Rock type RMR Rock Class Rock category

0+000m to 0+100m 100m Schist 45-55 III Fair

0+100m to 0+200m 100m Schist 25-35 IV Poor

0+200m to 0+707m 505m Gneiss 62-71 II Good

0+707m to 0+876 m 169m Gneiss 25-30 IV Poor

0+876m to 1+938m 1062m Gneiss 53-60 III Fair



Table 5.2: Rock mass classification along headrace tunnel based on Q value.

Chainage Length Rock type Q Rock Class Rock catagory

0+000m to 0+100m 100m Schist 5.20 III Fair

0+100m to 0+200m 100m Schist 3.75 IV Poor

0+200m to 0+707m 505m Gneiss 13.69 II Good

0+707m to 0+876 m 169m Gneiss 3.75 IV Poor

0+876m to 1+938m 1062m Gneiss 4.58 III Fair



5.5 Engineering Geological Condition of Inclined Shaft And Penstock Tunnel

An inclined (55º) shaft is proposed just after surge shaft which will have a finished

diameter of 4 m. The surface geological mapping from surge tank to the powerhouse

indicates that the area above the alignment is mostly covered by colluviums deposit. The

bedrock in the inclined shaft area is gneiss (Figure 5.13). The rock exposed in the inclined

shaft area is moderately weathered, medium- to thinly-banded and medium strong rock.

The same value of rock mass rating for surge tank area is also applied to the rock mass of

the inclined shaft area. Hence the rock mass of the inclined shaft area is categorized as

fair quality rock belonging to class III type. The trend of rock and joint system are similar

to that of surge tank area. The strike of the rock is nearly perpendicular to the alignment

of inclined shaft. The orientation of rock is favorable for the construction of incline shaft

and penstock.

5.6 Engineering Geological Condition of Powerhouse Site

The underground powerhouse of size 42.6 m (L) × 14 m (B) × 30.2 m (H) is proposed

about 500 m upstream from confluence of the Andheri Khola and the Trilsuli River on the

right bank of the Trisuli River at Pairegaun. The surface of the powerhouse area is

covered by colluviums (Figure 5.14 and 5.15). The colluviums consist of the boulder and

cobbles of gneiss and schist. The rock mass condition of powerhouse area has been

extrapolated from the mapping in the river section and was projected to the powerhouse

in order to produce the required geological condition. The majority of joints are closely to

widely spaced, tight to open with rough, irregular, planar, smooth and occasionally

undulating surface. The persistency of joint is generally continuous. The joint surfaces are

mainly fresh with occasionally iron staining and silty coating. The RMR and Q values of

the rock mass around powerhouse site range from 61-68 and 5.83-10 respectively which

place the rock in category II to III and described as good to fair rock. The statistical

analysis of discontinuities measured around powerhouse site (Figure 5.16) shows the

three set of major joints which are F: 224º/21º, J1: 109º/44º and J2: 018º/78º. Three

wedges are seen out of which W1 is in daylight condition. W2 and W3 seem to be stable

because their dip is less than friction angle.

Figure 5.15: Photograph of powerhouse site near Paire guan.


near Simle.

The proposed underground powerhouse site has been investigated by drill hole DP-3 and

seismic profiles SLP-6 to SLP-10 (Annex II). The drill hole DP-3 has shown the bedrock

at 28.90 m depth and seismic profile has shown similar result. The rock type is gneiss

which is slightly weathered, hard and compact and consist mainly three set of joints. The

overall core recovery and overall RQD are 96% and 51% respectively. The drilled core

showed mostly fresh joint surface with occasional iron staining and clay filling.

54

The orientation of the powerhouse is N67º with its longitudinal axis nearly perpendicular

to the strike of the foliation which is a favorable condition with respect to the stability of

underground powerhouse excavation.

5.7 Engineering Geological Condition Of Tailrace Tunnel

About 50% of the tailrace tunnel is expected to be located in gneiss rock, which is similar

to the rock at powerhouse site. The surface mapping showed the good to fair quality rock.

The rest 50% of the tailrace tunnel is expected to excavate in overburden deposit which

consist mainly of alluvium. The alluvial deposit in tailrace tunnel consist mainly of

rounded to sub-rounded boulder and gravel of schist, quartzite and gneiss mixed in sandy-

silty matrix.

The tailrace tunnel alignment has been investigated by drill hole DP-2 which is drilled up

to 37.80 m. The bedrock has not been encountered in the drill hole. Hence the bedrock

level is expected to be at more than 40.00 m depth and similar result has also been by

seismic profiles SLP-6, SLP-10, SLP-11, SLP-12, SLP-13 and SLP-14 (Annex II). The

tailrace tunnel is nearly parallel to the strike of foliation having 20º-30º dip amount which

gives fair tunneling condition for excavation.

55

CHAPTER SIX

GEOTECHNICAL STUDY OF THE

UNDERGROUND STRUCTURES Geotechnical study of the proposed headrace tunnel include establishment of geotechnical

parameters in order to know the interaction between the existing ground condition and the

proposed structure. Data required for geo-technical studies of head race tunnel are

acquired from geological and engineering geological mapping, core drilling and empirical

techniques. Since all necessary parameters for geotechnical studies are not available at the

present level of study, those parameters which are the most essence for the geotechnical

studies are determined using empirical relationships. Geotechnical studies include

preliminary stress analysis, underground wedge stability analysis and rock support design

along the headrace tunnel. However, at this stage of study all the data required to carry

out the analysis can not be obtained. Therefore, the following assumptions were

considered during the geotechnical design of the underground structure of the project.

The in situ stress and elastic parameter of the rock are difficult to acquire at this

stage of study level. So the major principal stress (σ1) is assumed to be equal to

the vertical stress due to overburden and minor principal stress (σ3) is assumes to

be 0.58 times the vertical stress with addition of tectonic stress component of 1

MPa and the intermediate principal or out of plane stress (σ2) is assumed to the

sum of the minor principal stresses and the tectonic stress component of 1MPa.

It is assumed that the major principal stress (σ1) is oriented in vertical direction

and the minor principal stress (σ3) is oriented in the horizontal direction

perpendicular to tunnel and cavern axis and intermediate stress (σ2) is oriented in

the horizontal direction parallel to the tunnel alignment.

Elastic and Plastic parameter are taken form empirical relation proposed by Hoek

et al. (1995)

56

6.1 STRESS ANALYSIS ALONG UNDERGROUND STRUCTURES

On the present study an attempt is made for the analysis of stress condition produced by

overburden rock body along the headrace tunnel using RMR, GSI (Geological strength

index) and Q which are extracted form surface mapping and other values obtained fro

different empirical methods. This includes determination of in situ stress deformation

modulus, elastic and plastic behavior and failure criteria.

6.1.1 Estimation of In situ Deformation Modulus

In situ deformation modulus (Em) of rock is an important parameter in any form of

numerical analysis related to stability of rock masses. But this parameter is difficult and

expensive to determine in field and are not generally determined at this study level.

However, this parameter can be determined empirically using the following relations.

These relations use rock mass classification RMR and Q which are calculated on the

present study.

Em = 2RMR-100 for RMR>55 (Bieniawski, 1978)……………..… (6.1)

Em = 10(RMR-10)/40 (Serafim and Pereira, 1983)……………………. (6.2)

Em = 25 log10Q (Girmstad and Barton, 1993)……………………...(6.3)

Where,

Em is in situ deformation modulus of rock mass in GPa

RMR is rock mass rating, and

Q is tunneling quality index.

In situ deformation (Em) along the proposed headrace tunnel is calculated using RMR and

Q form the equations 6.1, 6.2 and 6.3. Average in situ deformation is obtained and

presented in Table 6.1

57

Table 6.1: Estimation of in situ deformation of rock along underground structures

Structure Chainage Rock type RMR Q

Em= 2RMR-100

(GPa)

Em= 10(RMR-

10)/40

(GPa)

Em= 25log10Q

(GPa)

Average Em (GPa)

Headrace tunnel

0+000m - 0+100 m Schist 53 5.20 11.88 17.90 14.89

0+100m – 0+200m Schist 35 3.75 4.23 14.35 9.28

0+200m – 0+707m Gneiss 66 13.69 32 28.41 30.20

0+707m – 0+876m Gneiss 35 3.75 4.23 14.35 9.28

0+876m -1+938m Gneiss 53 4.58 11.89 16.52 14.20

1+938m – 2+600m Gneiss 65 9.90 30 24.89 27.44

2+600m – 4+142m Gneiss 63 7.48 26 21.85 23.92

Surge shaft Gneiss 63 5.83 26 19.14 22.57

Inclined shaft Gneiss 63 5.83 26 19.14 22.57

Power house Gneiss 65 7.90 30 22.44 26.22

Tailrace tunnel Gneiss 65 7.90 30 22.44 26.22

6.1.2 In situ Stress Analysis

Basically, the governing parameter for the stability of the rock inside the tunnel is

orientation of joints, its separation and pressure caused by the overburden or rock cover.

In order to avoid the hydraulic fracturing of rock with the consequent opening of existing

joints, the minor principal component of in situ stresses should be higher than an internal

hydrostatic pressure in the tunnel. At this level of study it is expensive to carry out the test

required to measure in situ stress. Therefore, an empirical method is used here for the

evaluation of in situ stress along the headrace tunnel.

The vertical stress acting on a tunnel is estimated from the simple relationship

σv = γz …………………………………………………………….. (6.4)

Where,

σv is vertical stress

γ is unit weight of overlying rock body, and

z is depth below a surface

58

Horizontal stress acting on the tunnel at depth z below a surface can be estimated as,

σh = k σv…………………………………………………………….(6.5)

Where,

σh is horizontal stress, and

k is ratio of horizontal to vertical stress

Sheory (1994) has given an empirical equation to estimate the value of horizontal to

vertical stress ration (k) as

k = 0.25+7Em(0.001+1/z)……………………………………..…….(6.6)

Where,

z is depth below a surface in meter

Em is average deformation modulus of upper part of earth crust’s

measured in horizontal direction in GPa

Vertical and horizontal stress as well as horizontal to vertical stress ratio (k) along the

headrace tunnel is calculated and presented in Table 6.2. Overburden from the crown of

the tunnel to the surface level is used as the maximum rock cover (z) ignoring the depth

of residual/colluviums cover above the bed rock though residual/colluviums cover varied

at different level along the tunnel alignment. Unit weight of rock (γ) is assumed as 0.027

MN/m3 and in situ deformation modulus (Em) is taken from Table 6.1.

6.1.3 Determination of Elastic and Plastic Behavior of Rock

Different stress parameters like vertical stress, maximum tangential boundary stress, in

situ deformation modulus and ratio of horizontal to vertical stress are used to find out

elastic and plastic behavior of rock. The ratio of maximum tangential boundary stress to

the unconfined compressive stress of rock mass is referred as damage index (Di). The

damage index is thus given by a relation,

Di = σmax/σc………………………………………………………...(6.7)

Where,

59

σmax is maximum tangential boundary stress, and

σc is unconfined compressive strength

If Di ≤0.4, rock behaves as elastic and if Di>0.4, rock behaves as plastic.

Maximum tangential boundary stress (σmax) is given by the Kirsch equation,

σmax = σv (3k-1) (Hoek and Brown, 1980)………………………..(6.8)

Where,

k is horizontal to vertical stress ratio, and

σv is vertical stress

Table 6.2: Estimation of in situ horizontal and vertical stress along underground

structures

Structure Chainage Rock type

z (m)

γ (MN/m3)

Em (GPa)

σv (MPa)

k

σh (MPa)

Headrace tunnel

0+000m - 0+100 m Schist 75 0.027 14.89 2.025 1.744 3.531

0+100m – 0+200m Schist 142.5 0.027 9.28 3.847 0.771 2.966

0+200m – 0+707m Gneiss 225 0.027 30.20 6.075 1.401 8.511

0+707m – 0+876m Gneiss 202.5 0.027 9.28 5.467 0.636 3.476

0+876m -1+938m Gneiss 210 0.027 14.20 5.670 0.823 4.665

1+938m – 2+600m Gneiss 165 0.027 27.44 4.455 1.606 7.156

2+600m – 4+142m Gneiss 300 0.027 23.92 8.100 0.976 7.902

Surge shaft Gneiss 75.75 0.027 22.57 2.045 2.493 5.100

Inclined shaft Gneiss 277.5 0.027 22.57 7.492 0.977 7.323

Power house Gneiss 202.5 0.027 26.22 5.467 1.339 7.326

Tailrace tunnel Gneiss 150 0.027 26.22 4.050 1.657 6.711

Damage index for the headrace tunnel is estimated and presented in Table 6.3. Vertical

and horizontal to vertical stress ratio is taken from Table 6.2 and unconfined compressive

strength (UCS) obtained by field and laboratory (Annex I) is used for determination of

damage index.

60

Table 6.3: Damage index of rock mass along underground structures.

Structure Chainage Rock type σv (MPa) k σc

(MPa) σmax

(MPa) Di

Headrace tunnel

0+000m - 0+100 m Schist 2.025 1.744 35 8.569 0.245

0+100m – 0+200m Schist 3.847 0.771 35 5.0511 0.144

0+200m – 0+707m Gneiss 6.075 1.401 100 19.4582 0.195

0+707m – 0+876m Gneiss 5.467 0.636 100 4.9640 0.049

0+876m -1+938m Gneiss 5.670 0.823 100 8.3292 0.083

1+938m – 2+600m Gneiss 4.455 1.606 100 17.0091 0.170

2+600m – 4+142m Gneiss 8.100 0.976 100 15.6168 0.156

Surge shaft Gneiss 2.045 2.493 100 13.2495 0.132

Inclined shaft Gneiss 7.492 0.977 100 14.4670 0.145

Power house Gneiss 5.467 1.339 100 16.4939 0.165

Tailrace tunnel Gneiss 4.050 1.657 100 16.0825 0.161

As mentioned earlier, for damage index Di≤0.4, rock mass behaves as an elastic condition

and no visible damage occurs. On the present study, all calculated values of Di along the

underground structures are <0.4. Hence, the rock mass behaves as an elastic behavior.

This concludes that there is no possibility of damage in the tunnel due to overburden rock

body.

6.1.4 Determination of Rock Mass Strength along the Headrace Tunnel.

The rock mass properties are assumed to be adequately characterized by the biaxial

failure criteria developed by Hoek and Brown. The most general form of Hoek-Brown

criterion which incorporates both original and modified form is given by the following

equation for both intact and fractured rock.

σ1 = σ3 + (mbσcσ3+sσc2)1/2…………………………………………... (6.9)

Where,

σ1 is major principal stress at failure

σ3 is minor principal stresses applied to the specimen

61

σc is uniaxial compressive strength of intact rock material in the

specimen

mb and s are material constants which depend upon properties of

rock and upon an extent to which it has been broken before

subjected to stresses σ1 and σ3

Uniaxial compressive strength (σcs) of a specimen is given by substituting σ3 = 0 in above

equation, giving following equation:

σcs = (s σc2)1/2…………………….. ……………………….. (6.10)

For the intact rock, σcs = σc and s = 1, mb = mi. For the previously broken rock, s <1 and

the strength at zero confining pressure is given by above equation.

The uniaxial tensile strength of the specimen (σt) is given by substituting σ1 = 0 in the Eq.

6.9 and by solving the resulting quadratic equation for σ3

σt = ½ σc (mb – (mb2

+ 4s) ½) ……………………………….(6.11)

The strength parameters, m and s for the intact and the fractured rock are as follows.

Intact rock s = 1

Very fractured rock s = 0

Good quality rock mi = 25

Weak rock mi = 0

Values of mb and s used in analysis is determined from the following equation and RMR

is determined according to Bieniawski (1989)

For GSI> 25 (undisturbed rock masses)

mb/mi = exp (GIS-100)/28……………………………………… (6.12)

s = exp (GSI-100)/9 ……………………….................................(6.13)

Where,

62

GSI is geological strength index

The relation between GSI and RMR is given by an equation,

GSI = RMR89 – 5………………….………. ………………(6.14)

Where,

RMR89 is rock mass classification proposed by Bieniawski

(1989)

Values of constant mi for the intact rock are given in Table 6.4

In order to determine rock mass strength parameters, mb and s, GSI calculated and

tabulated on Table 6.4 is taken. The value of mi is taken from Table 6.5. Thus determined

strength parameter is tabulated on Table 6.6. The Mohr-Coulomb failure criteria is listed

in Annex V.

Table 6.4: Determination of rock mass strength parameter, mb and s


Maximum rock cover, z (m) GSI σc

(MPa) mi mb s

Headrace tunnel

0+000m - 0+100 m Schist 75 48 35 12 1.8734 0.0031

0+100m - 0+200m Schist 142.5 30 35 12 0.9850 0.0042

0+200m -0+707m Gneiss 225 61 100 28 6.9542 0.1312

0+707m -0+876m Gneiss 202.5 30 100 28 2.2984 0.0042

0+876m -1+938m Gneiss 210 48 100 28 4.3713 0.0031

1+938m -2+600m Gneiss 165 60 100 28 6.7102 0.0117

2+600m -4+142m Gneiss 300 58 100 28 6.2476 0.0094

Surge shaft Gneiss 75.75 58 100 28 6.2476 0.0094

Inclined shaft Gneiss 277.5 58 100 28 6.2476 0.0094

Power house Gneiss 202.5 60 100 28 6.7102 0.0117

Tailrace tunnel Gneiss 150 60 100 28 6.7102 0.0117

63

Table 6.5: Values of mi for intact rock (Marrinos and Hoek, 2001)

Table 6.6: Analysis of rock strength using Roclab.

Structure Headrace tunnel Surge shaft

Inclined shaft

Power house

Tailrace tunnel

Chainage 0+000m – 0+100m

0+100m – 0+200m

0+200m – 0+707m

0+707m – 0+876m

0+876m – 1+938m

1+938m – 2+600m

2+600m – 4+142m

Rock type Schist Schist Gneiss Gneiss Gneiss Gneiss Gneiss Gneiss Gneiss Gneiss gneiss

Hoek-Brown classification

Intact uniaxial compressive strength (σc, MPa) 35 35 100 100 100 100 100 100 100 100 100

GSI 48 30 61 30 48 60 58 58 58 60 60 mi 12 12 28 28 28 28 28 28 28 28 28

Disturbance factor (D) 0 0 0 0 0 0 0 0 0 0 0

0 mb 1.873 0.985 6.954 2.298 4.371 6.710 6.248 6.248 6.248 6.710 6.710

s 0.0031 0.0004 0.0131 0.0004 0.0031 0.0117 0.0094 0.0094 0.0094 0.0117 0.0117 a 0.507 0.522 0.503 0.522 0.507 0.503 0.503 0.503 0.503 0.503 0.503

Failure envelope range

Application Tunnel Tunnel Tunnel Tunnel Tunnel Tunnel Tunnel Tunnel Tunnel Tunnel Tunnel

σ3max (MPa) 1.0195 1.8182 3.1762 2.7663 2.9304 2.3701 4.0028 1.1373 3.8542 2.8732 2.1670 Unit weight (MN/m3) 0.027 0.0027 0.027 0.027 0.027 0.027 0.027 0.027 0.027 0.027 0.027

Overburden (m) 75 142.5 225 202.5 210 165 300 75.75 277.5 202.5 150

Mohr-Coulomb fit

Cohesion (MPa) 0.469 0.458 2.110 1.071 1.540 1.758 2.262 1.169 2.210 1.949 1.679

Friction angle 49.02º 38.58º 58.40º 51.00º 55.83º 60.03º 56.15º 63.84º 56.41º 58.82º 60.58º

Rock mass parameters

Tensile strength (MPa) -0.058 -0.015 -0.189 -0.018 -0.071 -0.175 -0.151 -0.151 -0.151 -0.175 -0.175

Uniaxial compressive strength (MPa) 1.875 0.602 11.325 1.720 5.356 10.701 9.550 9.550 9.550 10.701 10.701

Global strength (MPa) 6.375 4.215 35.857 18.678 27.611 35.135 33.744 33.744 33.744 35.135 35.135 Deformation modulus

(MPa) 5272.71 1870.83 18836.49 3162.28 8912.51 17782.79 15848.93 15848.93 15848.93 17782.79 17782.79

6.2 UNDERGROUND WEDGE STABILITY ANALYSIS

The stability of the underground wedges that are likely to form around the headrace

tunnel and powerhouse cavern is carried out to determine the size of the wedges, their

mode of failure and factor of safety. Unwedge 3.005 software is used for the analysis. The

unit weight of the rock and friction angle is assumed to be 2.7 tones/m3 and 35º

respectively for analysis. The shear strength of the rock is considered to be zero.

The major three sets of joints with azimuth of tunnel alignment between Ch 0+000m to

Ch 0+876 m is plotted in streonet (Figure 6.1). The major joints and tunnel alignment

forms the different wedges at roof, sidewall and floor of the tunnel as shown in Figure

6.2.

Figure 6.1: Stereoplot of major joint sets within Ch 0+000 m to 0+876 m

66

Figure 6.2: Wedges expected in tunnel in between Ch 0+000 m to 0+876 m

The analysis shows mainly two critical wedges having factor safety less than or equal to

one. Wedge No. 6 formed at roof of the tunnel has weight of 3.894 tonnes and apex

height 0.70 m with factor of safety 0.255. This wedge will slide on J2 (70º/020º). Wedge

No. 7 formed at sidewall (right) has weight of 30.659 and apex height 3.13 with factor of

safety 1.054. This wedge will slide along the intersection of J2 (70º/020º) and J3

(41º/079º). To stabilize these wedges bolting and shotcrete is required.

To stabilize the Wedge No 6 formed at the roof of the tunnel, shotcrete of 10 cm thick

having shear strength of 25 tonnes/m2 with unit weight of 2.7 tonnes/m3 and mechanically

anchored type 2 m long rock bolt having tensile, plate and anchor capacity of 10 tonnes

are applied at normal to boundary in 2 m × 2 m spacing. After installation of support, the

factor of safety will become 7.546 (Figure 6.3)

Mechanically anchored type 4 m long rock bolt having tensile, plate and anchor capacity

of 10 tonnes are applied at normal to boundary in 2.5 m × 2.5 m spacing to stabilize the

Wedge No 7. The factor of safety increases to 3.719 after installation of support (Figure

6.4).

67

Figure 6.3: Support applied to stabilize Wedge No 6


The major three sets of joints with azimuth of tunnel alignment between Ch 0+876 m to Ch 1+938 m is plotted in streonet (Figure 6.5).

68

Figure 6.5: Stereoplot of major joint set within Ch 0+876 m to Ch 1+938 m

The major joints and tunnel alignment forms the different wedges at roof, sidewall and

floor of the tunnel as shown in Figure 6.6.

Figure 6.6: Wedges expected in the tunnel within Ch 0+876 m to Ch 1+938 m

69

Figure 6.7: Support applied to stabilize Wedge No 6 and 8


The analysis shows that there are three critical wedges having factor safety less than one.

Wedge No. 6 formed at side wall (left) of the tunnel has weight of 0.298 tonnes and apex

height 0.36 m with factor of safety 0.404. This wedge will slide on J2 (60º/024º). Wedge

70

No. 7 formed at sidewall (right) has weight of 6.959 tonnes and apex height 3.13 with

factor of safety 0.241. This wedge will slide on J3 (71º/099º). Wedge No 8 formed at

crown of the tunnel has weight of 0.088 tonnes and apex height 0.15 with factor of safety

0.000. This wedge will freely fall under the influence of gravity. To stabilize these

wedges bolting and shotcrete is required.

To stabilize the Wedge No 6 and Wedge No 8, shotcrete of 10 cm thick having shear

strength of 25 tonnes/m2 with unit weight of 2.7 tonnes/m3 and mechanically anchored

type 2 m long rock bolt having tensile, plate and anchor capacity of 10 tonnes are applied

at normal to boundary in 2 m × 2 m spacing. After installation of support, the factor of

safety of Wedge No 6 and Wedge No 8 will become 22.652 and 53.891 respectively

(Figure 6.7).

Shotcrete of 10 cm thick having shear strength of 25 tonnes/m2 with unit weight of 2.7

tonnes/m3 and mechanically anchored type 2 m long rock bolt having tensile, plate and

anchor capacity of 10 tonnes are applied at normal to boundary in 2 m × 2 m spacing to

stabilize the Wedge No 7. The factor of safety increases to 8.323 after installation of

support (Figure 6.8).

The major three sets of joints with azimuth of tunnel alignment between Ch 1+938 m to


form the different wedges at roof, sidewall and floor of the tunnel as shown in Figure

6.10.

The analysis shows three critical wedges having factor safety less than one. Wedge No. 5

formed at roof of the tunnel has weight of 19.967 tonnes and apex height 1.29 m with

factor of safety 0.509. This wedge will slide on J2 (54º/130º). Wedge No. 7 formed at

roof has weight of 0.955 tonnes and apex height 0.41 m with factor of safety 0.136. This

wedge will slide on J3 (79º/183º). Wedge No 8 formed at roof has weight of 0.009 tonnes

with factor of safety 0.00. This wedge will fall under the influence of gravity. To stabilize

these wedges bolting and shotcrete is required.

71

Figure 6.9: Stereoplot of major joints within Ch 1+938 m to 2+600 m

Figure 6.10: Wedges expected in tunnel within Ch 1+938 m to 2+600 m

To stabilize the Wedge No 7 and 8, shotcrete of 10 cm thick having shear strength of 25

tonnes/m2 with unit weight of 2.7 tonnes/m3 and mechanically anchored type 2 m long

rock bolt having tensile, plate and anchor capacity of 10 tonnes are applied at normal to

72

boundary in 2.5 m × 2.5 m spacing. After installation of support, the factor of safety of

Wedge No 7 and 8 will become 17.96 and 123.56 respectively (Figure 6.11).


tonnes/m3 and mechanically anchored type 2.5 m long rock bolt having tensile, plate and

anchor capacity of 10 tonnes are applied at normal to boundary in 2 m × 2 m spacing to

stabilize the Wedge No 5. The factor of safety increases to 4.044 after installation of

support (Figure 6.12).

Figure 6.11: Support applied to stabilize Wedge No 7 and 8

73

Figure 6.12: Support applied for Wedge No 5

Figure 6.13: Stereplot of major joint sets within Ch 2+600 m to Ch 4+142 m

The major three sets of joints with azimuth of tunnel alignment between Ch 2+600 m to


74

forms the different wedges at roof, sidewall and floor of the tunnel as shown in Figure

6.14.

Figure 6.14: Wedges expected in tunnel within Ch 2+600 m to Ch 4+142 m

The analysis shows two critical wedges having factor safety less one. Wedge No. 6 has

weight of 14.106 tonnes and apex height 2.03 m with factor of safety 0.725. This wedge

will slide on J2 (44º/109º). Wedge No. 8 has weight of 0.924 tonnes and apex height 0.45

m with factor of safety 0.00. This wedge fall under the influence of gravity. To stabilize

these wedges bolting and shotcrete is required.

To stabilize the Wedge No 6, shotcrete of 10 cm thick having shear strength of 25

tonnes/m2 with unit weight of 2.7 tonnes/m3 and mechanically anchored type 2.5 m long

rock bolt having tensile, plate and anchor capacity of 10 tonnes are applied at normal to

boundary in 2.5 m × 2.5 m spacing. After installation of support, the factor of safety will

become 6.155 (Figure 6.15)

Mechanically anchored type 2 m long rock bolt having tensile, plate and anchor capacity

of 10 tonnes are applied at normal to boundary in 2.5 m × 2.5 m spacing to stabilize the

Wedge No 8. The factor of safety increases to 14.206 after installation of support (Figure

6.16).

75



The major three sets of joints with azimuth of tunnel alignment of powerhouse cavern is

plotted in streonet (Figure 6.17). The major joints and tunnel alignment forms the

different wedges at roof, sidewall and floor of the powerhouse cavern as shown in Figure

6.18

76

Figure 6.17: Stereoplot of major joint set in powerhouse cavern

Figure 6.18: Wedge expected in powerhouse cavern.

The analysis shows three critical wedges having factor safety less one. Wedge No. 6 has

weight of 3190.535 tonnes and apex height 12.56 m with factor of safety 0.725. This

wedge will slide on J2 (44º/109º). Wedge No. 7 has weight of 2.198 tonnes and apex

77

height 0.69 with factor of safety 0.149. This wedge will slide on joint J3 (78º/018º).

Wedge No 8 has weight of 0.530 tonnes and apex height 0.35 m with factor of safety

0.00. This wedge will freely fall under the influence of gravity. To stabilize these wedges

bolting and shotcrete is required.

To stabilize the Wedge No 6, shotcrete of 15 cm thick having shear strength of 150

tonnes/m2 with unit weight of 2.7 tonnes/m3 and mechanically anchored type 8 m long

rock bolt having tensile, plate and anchor capacity of 10 tonnes and shear strength of 200

tonnes are applied at 2 m × 2 m spacing. After installation of support, the factor of safety

will become 3.296 (Figure 6.19).



tonnes/m3 and mechanically anchored type 2.5 m long rock bolt having tensile, plate and

anchor capacity of 10 tonnes and shear strength 200 tonnes are applied at normal to

boundary in 2.5 m × 2.5 m spacing to stabilize the Wedge No 7 and 8. The factor of

safety increases to 82.853 and 153.461 respectively after installation of support (Figure

6.20).

78

Figure 6.20: Support applied for Wedge No 7 and 8

79

6.3 ROCK SUPPORT DESIGN

The proposed rock support design is based on today's common practice on tunnel

construction. Rock support is designed by a combination of direct observation, analytical

and numerical methods. Design is mainly based on classification of rock mass quality

along the tunnel. The input is provided by engineering geological surface mapping, field

investigations and laboratory testing of rock samples. RQD, RMR and Q and empirical

rules are adopted for the present design.

Combinations of rock bolts, fibre-reinforced and steel mesh-reinforced shotcrete and cast

concrete lining can be used as dictated by rock mass quality encountered under

excavation. A recommended principle for rock support is to require equal quality for the

initial support as for the permanent support. This requirement intends to incorporate the

initial support as part of the total requirement for permanent support. Experience indicates

that 50%-80% of the total required support is performed successively during excavation

and remaining 20%-50% later, thus leaving a relatively small volume of work to be

executed after excavation is finished. Adequate support is assessed based on relevant

properties of rock mass and support materials (shotcrete, rock bolts etc). The support is

installed and monitored as required and if necessary strengthened by an additional

support.

During tunnel excavation, types and quantities of rock support defined in the present

design can used as a menu for detailed design on construction phase. Detailed design is

adapted to in situ geological conditions either as registered at excavation front or as

indicated by investigations ahead of face on logging of geological conditions. Additional

information for the design is obtained by testing of materials sampled from the tunnel and

from direct registration of in situ stress conditions and behavior of proposed support; i.e.

by measuring deformation (convergence) of the cross section. Thus, the type and extent

of support needed is finalized only during the excavation. Since, the support system based

on the RMR system seems more reliable than based other system in present study, the

support recommended based on RMR system is suggested to be adopted as the menu for

the further investigation on the support system.

80

6.3.1 Rock Support Design Based On Rock Quality Designation (RQD)

In past centuries, ground support was always selected empirically. The miners estimated,

based on his experience, what timber was required, and if the timbering failed it was

rebuilt. Written rules for selecting ground support were first formulated by Terzaghi

(1946). Deer et al. (1970) correlated Terzaghi’s rock loads with approximate RQD values

and separate ground support recommendations for tunnels excavated conventionally and

for tunnels excavated by tunnel boring machine (Table 6.7). On the present study an

attempt is made to design the rock support based on RQD. Estimation of rock support

based on RQD is presented on Table 6.8.

Table 6.7: Support recommendations for Tunnels in Rock (6 m to 12 m dia.) based on

RQD (after Deere et al. 1970)

Rock quality Tunneling method

Alternative support systems

Steel sets3 Rockbolts3 shotcrete

Excellent1 (RQD>90)

Boring machine

None to occasional light set, rock load (0.0-0.2)B

None to occasional

None to occasional local application

Conventional None to occasional light set, rock load (0.0-0.3)B

None to occasional

None to occasional local application 2 to 3 in

Good1 (75<RQD<90)

Boring machine

Occassional light sets to pattern on 5- to 6-ft center. Rock load (0.0 to 0.4)B

Occasional to pattern on 5- to 6-ft center.

None to occasional local application 2 to 3 in.

Conventional Light sets 5- to 6- ft center. Rock load (0.3 to 0.6)B

Pattern, 5- to 6-ft centers

Occasional local application 2 to 3 in.

Fair (50<RQD<75)

Boring machine

Light to medium sets, 5- to 6–ft center. Rock load (0.4 to 1.0)B

Pattern, 4- to 6-ft center 2- to 4-in. crown

Conventional Light to medium sets, 4- to 5-ft center. Rock load (0.6 to 1.3)B

Pattern, 3- to 5-ft center

4-in. or more crown and sides.

Poor2 (25<RQD<50

Boring machine

Medium circular sets on 3- to 4-ft center. Rock load (1.0 to 1.6)B


4- to 6-in. on crown and sides. Combine with bolts.

Conventional Medium to heavy circular sets on 2- to 4-ft center. Rock load (1.3 to 2.0)B


6-in. or more on crown and sides. Combine with bolts

Very poor3 (RQD<25, excluding squeezing or swelling

ground)

Boring machine

Medium to heavy circular sets on 2-ft center. Rock load (1.6 to 2.2)B


6-in. or more on whole section. Combine with medium sets.

Conventional Heavy circular sets on 2-ft center. Rock load (1.6 to 2.2B)

Pattern, 3-ft center

6-in. or more on whole section. Combine with medium sets.

Very poor3 (RQD<25, squeezing or swelling ground)

Boring machine

Very heavy circular sets on 2-ft center. Rock load up to 250 ft.


6-in. or more on whole section. Combine with heavy sets

Conventional Very heavy circular sets on 2-ft center. Rock load up to 250 ft.


6-in. or more on whole section. Combine with heavy sets

Notes: 1 In good and excellent rock, the support requirement will be, in general, minimal but will be dependent upon joint geometry, tunnel diameter, and relative orientation of joints and tunnel. 2 Lagging requirements will usually be zero in excellent rock and will range from up to 25 percent in good tock to 100 percent in very poor rock.

81

3 Mesh requirements usually will be zero in excellent rock and will range from occasional mesh (or strips) in good rock to 100 percent mesh in very poor rock. 4 B = tunnel width

Table 6.8: Estimation of rock support for underground structures based on RQD.


RQD (%)

Rock mass

quality

Proposed support system

Steel set Rock bolts Shotcrete

Headrace tunnel

0+000m - 0+100 m Schist 64 Fair Light to medium sets on 5

to 6ft center Pattern, 5 to 6ft

center Occasional local

application 2 to 3in.

0+100m - 0+200m Schist 45 Poor

Medium to heavy circular sets on 2- to 4-ft center. Rock load (1.3 to 2.0)B


6-in. or more on crown and sides.

Combine with bolts 0+200m -0+707m Gneiss 80 Good Light sets 5- to 6- ft center.

Rock load (0.3 to 0.6)B Pattern, 5- to 6-

ft centers Occasional local

application 2 to 3 in.

0+707m -0+876m Gneiss 45 Poor




Combine with bolts 0+876m -1+938m Gneiss 58 Fair Light to medium sets on 5



application 2 to 3in. 1+938m -2+600m Gneiss 73 Fair Light to medium sets on 5



application 2 to 3in. 2+600m -4+142m Gneiss 70 Fair Light to medium sets on 5

to 6ft centerPattern, 5 to 6ft



Surge shaft Gneiss 50 Poor




Combine with bolts

Inclined shaft Gneiss 50 Poor




Combine with bolts Power house Gneiss 55 Fair Light to medium sets on 5



application 2 to 3in. Tailrace tunnel Gneiss 55 Fair Light to medium sets on 5

to 6ft centerPattern, 5 to 6ft



6.3.2 Rock Support Design Based on Rock Mass Rating (RMR)

Bieniawski (1989) has proposed a guide for the choice of support for underground

excavation based on RMR. The support guideline for horseshoe shape tunnel having

diameter of 10m, vertical stress less than 25MPa and method of construction is drilling

and blasting is shown in Table 6.9. The support estimated for the underground structures

based on RMR is given in Table 6.10.

82

Table 6.9: Geomechanics classification guide for excavation and support in rock tunnels

after Bieniawski (1989)

Rock Mass Excavation

Support

Rockbolts (20mm dia. Fully bonded

Shotcrete Steel sets

Very Good Rock,

I RMR: 81-100

Full Face 3m advance Generally no support required except for occasional spot bolting

Good Rock II

RMR: 61-80

Full face 1.0-1.5 m advance. Complete support

20 m from face

Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh

50 mm in crown where required None

Fair Rock, III

RMR: 41-60

Top heading and bench 1.5-3m advance in heading. Commence support after

each blast. Complete support 10m from face.

Systematic bolts 4m long, spaced 1.5-2m in crown and walls with mesh in crown.

50-100mm in crown, 30 mm in side wall

None

Poor Rock, IV

RMR: 21-40

Top heading and bench, 1-1.5m advance in heading.

Install support concurrently with excavation 10m from

face

Systematic bolts 4-5m long, spaced 1-1.5m in crown and walls with wire mesh

100-150mm in crown, and 100mm in sides

Light ribs spaced 1.5m where required

Very Poor Rock

V, RMR; <20

Multiple drifts. 0.5-1.5m advance in top heading.

Install support concurrently with excavation. Shotcrete as soon as possible after

blasting

Systematic bolts 4-5m long spaced 1-1.5m in crown and walls with wire mesh. Bolts invert

150-200mm in crown 150mm on sides and 50mm on face

Medium to heavy ribs spaced 0.75 m with steel lagging and fore poling if required. Close invert

6.3.3 Rock Support Design Based On Tunneling Quality Index (Q)

In order to relate tunnel quality index, Q, to support requirement of an underground

excavation, Barton et al. (1974), defined a parameter, which is referred as equivalent

dimension (De) of excavation. This dimension is obtained by dividing span diameter span

or wall height of excavation by the quantity called excavation support ratio (ESR).

Hence,

De = RatioSupport Excavation

(m)height or diameter span, Excavation

The excavation support ratio is related to the use for which the excavation is intended and

the extent to which some degree of instability is acceptable. The ESR for different types

of tunnels are given by Barton et al. (1974) is given in Table 6.11.

83

Table 6.10: Estimation of excavation and support in underground structures based on

RMR.


RMR

Rock class Excavation

Support

Rock bolts Shotcrete Steel ribs

Headrace

tunnel

0+000m - 0+100 m Schist 53 Fair

(III)

Top heading and bench 1.5-3m advance in heading. Commence support after each blast. Complete support 10m from face.



None

0+100m – 0+200m Schist 35 Poor

(IV)

Top heading and bench, 1-1.5m advance in heading. Install support concurrently with excavation 10m from face




0+200m – 0+707m Gneiss 66 Good

(II)

Full face 1.0-1.5 m advance. Complete support 20 m from face


50 mm in crown where required

None

0+707m – 0+876m Gneiss 35 Poor

(IV)

Top heading and bench, 1-1.5m advance in heading. Install support concurrently with excavation 10m from face




0+876m -1+938m Gneiss 53 Fair

(III)

Top heading and bench 1.5-3m advance in heading. Commence support after each blast. Complete support 10m from face.



None


(II)




None


(II)




None

Surge

shaft Gneiss 63 Good (II)




None

Inclined

shaft Gneiss 63 Good (II)




None

Power

house Gneiss 65 Good (II)




None

Tailrace

tunnel Gneiss 65 Good (II)




None

84

Table 6.11: Value of the ESR for the different types of the excavation category (Barton et

al. 1974) Class Excavation category ESR

A Temporary mine opening 3-5

B Permanent mine opening, water tunnels for hydropower (excluding high pressure penstock) pilot tunnels, drafts and headings for large excavation.

1.6

C Storage rooms, water treatment plants, minor road and railway tunnels, surge chambers, access tunnels.

1.3

D Power stations, major road and railway tunnels, civil defenses chambers, portals, intersection. 1.0

E Underground nuclear power stations, railway stations, sports and publics facilities, factories 0.8

Equivalent dimension (De) is plotted against Q to define a number of support categories in

a chart (Barton et al. 1974). This chart has been updated by Grimstad and Barton (1993)

and has been reproduced as shown in Figure 6.1 which is used for rock support

assessment.

The headrace tunnel falls into a category for waters tunnel for hydropower and is assigned

an excavation support ratio (ESR) of 1.6 from Table 6.11. Hence, for an excavation span

(B) of 6.2m of headrace tunnel, equivalent dimension (De) is calculated as 3.875.

Estimated support along headrace tunnel is given in Table 6.12. Beside these support

design requirements, the values of excavation width, excavation support ratio and Q can

be used to determine rock bolt length and maximum unsupported span for excavation

using empirical relations.

Length of the rock bolt (L) can be estimated from excavation width (B) and excavation support ration (ESR) by the relation given below.

L = ( )ESR

B15.02 +

For the headrace tunnel, B = 6.2 m; minimum length of the rock bolt required, L = 1.83 m i.e. ~2 m.

Biron and Arigolu (1982) has adopted the relation between rock bolt and roof span and recommended the following.

For very strong roof: minimum recommended bolt length is 3-4 inches

For strong roof: bolt length = 1/3 of the roof span

For weak roof: ½ of roof span

85

Since the rock of study area is strong, the rock bolt given by Biron and Arigolu (1982) is 2m.

Maximum unsupported span for excavation can be estimated from Q and excavation support ratio (ESR) is given by relation

Maximum unsupported span = 2 ESR Q0.4

Estimation of maximum unsupported span for the headrace tunnel is given in Table 6.12.

The permanent roof support pressure is calculated by the relation below.

Proof = 3/1

32 −× Q

JrJn

Where,

Proof is permanent roof support pressure

Jn is joint set number

Jr is joint roughness number

Q is NGI tunneling quality index

Estimation of permanent roof support pressure is presented in Table 6.12

Figure 6.21: Estimated support categories based on the tunneling quality index (Q)

86

Table 6.12: Estimation of rock support based on Q

Structure Chainage Rock type Q Rock

class

Max. unsupported span (m)

Proof (Kg/cm2)

Support category

Support

Rock bolts Shotcrete

Headrace tunnel

0+000m - 0+100 m Schist 5.20 Fair

(III) 6.19 0.768 III

Systematic bolting, 2m long, 2m (center to center) spacing

100mm fibre-reinforced shotcrete at crown and wall

0+100m – 0+200m Schist 3.75 Poor

(IV) 5.43 0.856 IV

Systematic bolting, 2m long, 1.5m (center to center) spacing


0+200m – 0+707m Gneiss 13.69 Good

(II) 9.11 0.556 II

Spot bolting, 2m long, 2.5m (center to center) spacing


0+707m – 0+876m Gneiss 3.75 Poor

(IV) 5.43 0.856 IV

Systematic bolting, 2m long, 1.5m (center to center) spacing


0+876m -1+938m Gneiss 4.58 Fair

(III) 5.88 0.801 III



1+938m – 2+600m Gneiss 9.90 Fair

(III) 8.00 0.619 III



2+600m – 4+142m Gneiss 7.48 Fair

(III) 7.16 0.680 III



Surge shaft Gneiss 5.83 Fair

(III) 6.48 0.739 III



Inclined shaft Gneiss 5.83 Fair

(III) 6.48 0.739 III



Power house Gneiss 7.90 Fair

(III) 7.31 0.668 III



Tailrace tunnel Gneiss 7.90 Fair

(III) 7.31 0.668 III



6.3.4 Rock Support Design based on Empirical Design Recommendation According

to U.S. Corps of Engineers

A system of simple recommendations for rock bolt reinforcement design has been

formulated by U.S. Corps of Engineers (Table 6.13). This recommendation may be used

as a guide for minimum reinforcement required for tunnel. Rock support design

recommended for the headrace tunnel based on empirical design recommendation

according to U.S. corps of engineers is tabulated in Table 6.14. For recommendation of

87

rock bolt reinforcement value, bolt spacing is taken from Table 6.12. The thickness of

critical and potentially unstable rock block throughout the underground structures is

assumed as 1m.

Table 6.13: Recommendation of rock bolt reinforcement based on empirical design

recommendation to U.S. corps of engineers. Parameter Empirical rules

Min

imum

leng

th a

nd m

axim

um sp

acin

g

Minimum

length

Greatest of:

(a) 2 × bolt spacing

(b) 3 × thickness of critical and potentially unstable rock blocks 1

(c) For element above the spring line:

Spans < 6m =0.5 × span

Spans between 6m and 18m = interpolate between 3 and 4.5

Spans between 18m and 30m = 0.25 × span

(d) For elements below the spring line:

Height < 18m = same as (c) above

Height > 18m = 0.2 × height

Maximum

spacing

Least of:

(a) 0.5 × bolt length

(b) 1.5 × width of critical and potentially unstable rock blocks 1

(c) 2.0m 2

Minimum

spacing 0.9m to 1.2m

Notes: 1. Where joint spacing is close and span relatively large, the superposition of two reinforcement

patterns may be appropriate (e.g., long heavy elements on wide centers to support the span, and shorter, lighter bolts on closer centers to stabilize the surface against raveling).

2. Greater spacing than 2.0m makes attachment of surface support elements (e.g., weld-mesh or chain-link mesh) difficult.

88

Table 6.14: Estimation of rock support based on empirical design recommendation to U.S. corps of engineers


Recommendation for rock bolt reinforcement

Minimum length Maximum spacing Minimum spacing

Headrace tunnel

0+000m - 0+100 m Schist

Greatest of: (a) 2 × bolt spacing = 2 × 2m = 4m (b) 3 × thickness of potentially unstable block = 3 × 1m = 3m (c) for element above spring line = 3m (d) for element below spring line = 3m

Least of: (a) 0.5 × bolt length = 0.5 × 4m = 2m (b) 1.5 × width of critical and potentially unstable blocks = 1.5 × 1m = 1.5 (c) 2.0m

0.8m to 1.25m

0+100m – 0+200m Schist

Greatest of: (a) 2 × bolt spacing = 2 × 1.5m = 3m (b) 3 × thickness of potentially unstable block = 3 × 1m = 3m (c) for element above spring line = 3m (d) for element below spring line = 3m

Least of: (a) 0.5 × bolt length = 0.5 ×3m = 1.5m (b) 1.5 × width of critical and potentially unstable blocks = 1.5 × 1m = 1.5 (c) 2.0m

0.8m to 1.25m

0+200m – 0+707m Gneiss

Greatest of: (a) 2 × bolt spacing = 2 × 2.5m = 5m (b) 3 × thickness of potentially unstable block = (c) for element above spring line = 3m (d) for element below spring line = 3m

Least of: (a) 0.5 × bolt length = 0.5 × 5m = 2.5m (b) 1.5 × width of critical and potentially unstable blocks = 1.5 × 1 = 1.5 (c) 2.0m

0.8m to 1.25m

0+707m – 0+876m Gneiss

Greatest of: (a) 2 × bolt spacing = 2 × 1.5m = 3m (b) 3 × thickness of potentially unstable block = 3 × 1m = 3m (c) for element above spring line = 3m (d) for element below spring line = 3m

Least of: (a) 0.5 × bolt length = 0.5 ×3m = 1.5m (b) 1.5 × width of critical and potentially unstable blocks = 1.5 × 1m = 1.5 (c) 2.0m

0.8m to 1.25m

0+876m -1+938m Gneiss



0.8m to 1.25m

1+938m – 2+600m Gneiss



0.8m to 1.25m

2+600m – 4+142m Gneiss



0.8m to 1.25m

Surge shaft Gneiss


Least of: (a) 0.5 × bolt length = 0.5 × 4m = 2m (b) 1.5 × width of critical and potentially unstable blocks = 1.5 × 1m = 1.5

0.8m to 1.25m

89

(c) 2.0m

Inclined shaft Gneiss



0.8m to 1.25m

Power house Gneiss



0.8m to 1.25m

Tailrace tunnel Gneiss



0.8m to 1.25m

On the basis of above empirical and numerical methods, the following rock supports are

proposed for underground structures of the present project.

Table 6.15 Rock support for Upper Trisuli – 3A HEP.

S.N. Structure Fibre reinforced shotcrete Rock bolt

Crown Sidewall Length spacing

A. Headrace tunnel

1. Poor Rocks (Rock Class – IV) 150 mm 150mm 2.5 m 1.5 m × 1.5 m

2. Fair Rocks (Rock Class – III) 100 mm 100 mm 2.5 m 2 m × 2 m

3. Good Rocks (Rock Class – II) 2.5 m 2.5 m × 2.5 m

B. Surge Shaft

4 Fair Rocks (Rock Class – IV) 150 mm 150 mm 8 m 2 m × 2 m

C. Inclined Shaft

5. Fair Rocks (Rock Class – IV) 100 mm 100 mm 2.5 m 2 m × 2 m

D. Powerhouse

6. Fair Rocks (Rock Class – IV) 150 mm 150 mm 8 m 2 m × 2 m

E. Tailrace Tunnel

7. Fair Rocks (Rock Class – IV) 100 mm 100 mm 2.5 m 2 m × 2 m

The poor rock may require concrete lining to be stable. In Poor Rock about 40% of

shotcrete may be replaced by concrete lining on the basis of ground condition

requirement. The rock support pattern has to be reassessed during the construction stage.

The actual support pattern can only be recommended during the construction time.

90

CHAPTER SEVEN

CONCLUSIONS

7.1 CONCLUSIONS

The Upper Trisuli - 3A Hydroelectirc Project (HEP) lies in Nuwakot and Rasuwa district,

Central Nepal. The headwork area lies in Rasuwa district and the powerhouse area lies in

Nuwakot district. All the hydraulic structures are proposed on right bank of the Trisuli

River. Based on geological, engineering geological and geotechnical studies and

interpretation, following conclusions are drawn.

Geologically, Upper Trisuli - 3A HEP lies in the Kuncha Group of Lesser Himalayan

Metasediments in Central Nepal. In the project area, it is comprised of two units;

Schist Unit and Gneiss Unit

Schist Unit consists of Light grey, medium- to thick-banded psammatic schist with

occassional bands of medium- to thick-banded pelitic schist and medium-banded

quartzite.

Gneiss unit consist of milky white, medium- to thick-banded augen gneiss with

occassional partings of light grey to greenish grey schist.

In general, attitude of foliation plane of rock with in project area varies from NNE-

SSW to NNW-SSE dipping 10º-30º towards NE-SW.

The Main Boundary Thrust (MBT) passes about 80 km south and Main Central Thrust

(MCT) passes about 25 km north of the project area.

The bedrock peak ground acceleration value of the project area is 250-300 gal and

estimated design coefficient is 0.13-0.15.

The rock around headwork site is poor to good in quality. Strike of the foliation plane

is almost perpendicular to flow direction of the Trisuli River and dipping upstream at

headwork site. So, the area is favorable for construction of dam.

Both bank of the Trisuli River at weir axis are represented by terrace deposit consisting

of both alluvium and colluviums. The bedrock is 50 m away from the river bank on

right abutment and about 150 m far on left abutment. The bedrock is not observed up

to the depth of 30 m at weir axis. So, the dam will be founded in alluvium deposit.

91

The desander basin lies in alluvial deposit comprises of cobbles and pebbles of schist

and gneiss with silty, sandy and clayey matrix. There is an older rock fall consisting of

boulders of schist and gneiss up to size of 5 m. So, special attention should be given to

this rock fall for future stability.

There is not any major active debris flow in the vicinity of dam site. The Dharni Khola

is a major gully at headwork site. It seems that it is capable of bringing huge amount of

boulders. So, it may be dreadful during monsoon.

The outcrop at intake portal is cliff of light grey, slightly weathered, medium strong

schist with thin intercalation of quartzite. The joints are close to moderately close

spaced, tight to moderately open aperture, fresh to slightly weathered, medium

persistence, planar to rough surface with coating of sandy and silty material.

Orientation of foliation is less favorable for excavation. So, the outcrop should be

treated well before excavation.

95% of the headrace tunnel will passes through augen gneiss and only 5% will passes

through schist.

The rock mass along headrace tunnel ranges from poor rock (IV) to good rock (II).

The good rock (II) will occupies about 10%, fair rock (III) will occupies about 80%

and poor rock (IV) will occupies about 10% of the headrace tunnel on the basis of Q

values.

A local anticline is expected in tunnel alignment whose limbs in general dip NE and

SW. Expect thin bands of shear zones, there are not any major structural disturbances

such as fault and major shear zones. No any deep landslides intersect the tunnel

alignment.

The attitude of foliation is almost perpendicular to tunnel alignment dipping to NE-

NW and SE-SW with dip 10º -30º. Though the orientation of foliation is in favorable

direction, the gentle dip of foliation may create some problems during excavation.

The rock mass at surge shaft, inclined shaft, powerhouse area and tailrace tunnel varies

from fair rock (III) to good rock (II). Orientation of foliation is favorable but the gentle

dip may create some problems during the excavation.

Average in situ deformation modulus (Em) along the underground structures ranges

from 9.28 to 30.20 GPa.

92

Vertical stress (σv) and horizontal stress (σh) as well as horizontal to vertical stress

ratio (k) along the underground structures ranges 2.024 – 8.100 MPa, 2.966 – 8.511

MPa and 0.771 – 2.493 respectively.

Damage index (Di) along underground structure is less than 0.4 (0.049 – 0.195).

Hence, the rock mass behaves as an elastic.

Support design for various underground structures based on different systems

suggests the combination of local to systematic bolting and reinforced shotcrete as per

requirement.

93

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development in the Himalaya. Precambrian Research, 74, pp. 35-55.

Upreti, B.N., 1999. An overview of the stratigraphy and tectonics of the Nepal Himalaya.

Journal of Asian Earth sciences, 17, pp. 577-606.

98

Upreti, B.N., Le Fort, P., 1999. Lesser Himalayan crystalline nappes of Nepal: problem of

their origin. In: Macfarlane, A., Quade, J., Sorkhabi, R. (Eds.), Geological Society

of America Special paper, 328, pp. 225-238.

U. S. Army Corps of Engineers 1997. Engineering and Design, Tunnels and Shafts in

Rock, Washington, DC 20314-1000.

ANNEXS

Annex I Calculation of Uniaxial Confined Strength of Schist using Schimdt

Hammer Value and Laboratory Test result of Core Samples of Gneiss.

Annex II General Description of Drill Holes, Summary of Constant Head

Permeability Test Results and Brief Description of Seismic Refraction Survey

Annex III Rock Mass Rating System (After Bieniawski, 1989)

Annex IV Classification of indivisual parameters used in the Quality Index Q (After Barton et al 1974)

Annex V Mohr-Coulomb failure criteria

Annex VI Rock Mass Rating (RMR) and Tunnel Quality Index (Q) along Headrace

Tunnel

ii

Annex I

Calculation of Uniaxial Confined Strength of Schist using Schimdt Hammer Value and

Laboratory Test result of Core Samples of Gneiss.

iii

Calculation of Uniaxial Confined Strength of Schist using Schimdt Hammer Value.

S.N. Schmidt Hammer

Value α UCS

(MPa) S. N Schmidt Hammer

Valueα

UCS (MPa)

Mean UCS (MPa)

1 38 +90 80 89 24 -90 23 35.15 2 40 -90 55 90 18 +90 26 3 34 -90 40 91 22 +90 34 4 42 +90 100 92 42 +90 100 5 40 -90 55 93 58 +90 230 6 32 +90 58 94 18 +90 26 7 24 +90 38 95 32 -90 35 8 22 +90 34 96 24 -90 23 9 42 -90 55 97 38 +90 80 10 28 +90 45 98 36 +90 7511 22 +90 34 99 29 +90 48 12 32 +90 58 100 28 +90 45 13 50 +90 150 101 43 +90 105 14 42 +90 100 102 52 -90 105 15 26 +90 42 103 44 +90 110 16 22 +90 34 104 38 +90 80 17 40 +90 90 105 44 +90 110 18 42 +90 100 106 48 +90 137 19 34 -90 40 107 58 -90 139 20 36 +90 75 108 50 -90 5821 32 +90 58 109 48 +90 137 22 36 +90 75 110 54 +90 195 23 30 +90 52 111 46 +90 125 24 32 +90 50 112 40 +90 100 25 28 +90 45 113 52 -90 105 26 26 +90 42 114 54 -90 115 27 32 +90 58 115 44 -90 6828 42 +90 100 116 54 +90 198 29 42 +90 100 117 38 +90 80 30 44 +90 110 118 18 +90 26 31 44 +90 110 119 46 +90 125 32 56 +90 208 120 30 +90 52 33 38 +90 80 121 50 +90 150 34 42 +90 100 122 20 +90 30 35 37 +90 77 123 24 +90 28 36 44 +90 110 124 40 +90 90 37 32 +90 58 125 42 +90 100 38 38 +90 80 126 22 +90 34 39 32 -90 35 127 22 +90 34 40 38 +90 80 128 30 +90 52 41 26 -90 25 129 46 +90 125 42 38 -90 48 130 48 +90 137 43 32 +90 58 131 26 +90 42 44 28 +90 45 132 42 -90 60 45 46 +90 125 133 46 +90 125 46 46 +90 125 134 39 +90 88 47 22 +90 34 135 47 -90 80 48 32 +90 58 136 54 +90 154 49 32 +90 58 137 32 +90 58 50 24 +90 38 138 48 +90 13751 34 +90 65 139 39 +90 88 52 28 +90 45 140 48 +90 137 53 24 +90 44 141 44 +90 110

iv

54 30 +90 52 142 54 +90 175 55 40 +90 90 143 44 +90 110 56 48 +90 137 144 46 +90 125 57 14 +90 21 145 54 +90 195 58 20 +90 30 146 30 -90 32 59 22 +90 34 147 38 -90 48 60 18 +90 26 148 42 +90 100 61 16 +90 24 149 38 -90 48 62 24 +90 38 150 47 -90 80 63 16 +90 24 151 54 +90 115 64 24 -90 23 152 34 -90 40 65 44 +90 68 153 42 -90 60 66 17 +90 25 154 52 -90 105 67 18 +90 26 155 42 -90 60 68 18 +90 26 156 43 -90 65 69 25 +90 40 157 20 +90 30 70 20 +90 30 158 40 +90 90 71 16 +90 24 159 26 +90 42 72 35 +90 68 160 38 -90 48 73 15 +90 23 161 40 -90 55 74 18 +90 26 162 3 -90 3875 22 +90 34 163 30 +90 52 76 22 +90 34 164 28 +90 45 77 34 +90 65 165 22 +90 34 78 36 +90 75 166 26 +90 42 79 34 +90 65 167 20 +90 30 80 34 +90 65 168 28 +90 45 81 38 +90 80 169 28 +90 45 82 34 +90 65 170 36 -90 44 83 32 +90 58 171 24 +90 38 84 42 +90 100 172 24 +90 3885 54 +90 195 173 42 -90 60 86 47 +90 135 174 44 -90 68 87 47 +90 135 175 40 +90 90 88 39 +90 88 176 30 +90 52

Laboratory Test Result for Core Samples of Gneiss (NEA)

Depth m Location Specific Gravity Absorption % Unit Weight

gm/cm3

Uniaxial Compressive

Strength kg/cm2

Young Modulus Eav

N/mm2 31.60-32.00

Powerhouse

2.645 0.5 2.6 380.1 2941.4 46.33-46.90 2.66 0.3 2.6 930.8 10962.6 59.00-59.55 2.65 0.3 2.6 504.3 4845.9 69.55-70.00 2.70 0.4 2.6 1006.2 6235.8 26.00-26.50

Surge shaft

2.66 0.3 2.6 880.4 6695.0 35.23-35.70 2.67 0.3 2.6 955.9 6766.2 46.00-46.70 2.68 0.3 2.6 883.4 18866.3 58.00-58.65 2.64 0.4 2.6 1257.1 16913.4

v

Annex II

General Description of Drill Holes, Summary of Constant Head Permeability Test Results

and Brief Description of Seismic Refraction Survey

vi

General Description of Drill Holes (NEA)

Drill Hole No. Drilling Machine Inclination & Direction Location Length (m)

DH-1 Tone UD-5 Vertical Weir Axis (L/B) 30.20 DH-2 Tone UD-5 Vertical Weir Axis (R/B) 35.00 DH-3 Tone UD-5 Vertical Desander (R/B) 25.20 DP-3 Acker ‘Ace’ Vertical Underground Powerhouse 70.00 DP-4 Acker ‘Ace’ Vertical Surge Shaft 60.20

Summary of Constant Head Permeability Test Results (NEA)

Drill Hole No. Location Test Depth (m) Permeability Value cm/s

DH-1 Weir Axis (L/B)

17.05 1.01×10-1 24.30 6.80×10-3 27.75 5.00×10-3

DH-2 Weir Axis (R/B)

12.20 8.20×10-2 18.60 6.82×10-2 25.00 3.66×10-3 32.00 1.58×10-3

DH-3 Desander

5.20 2.22×10-2 10.20 4.62×10-2 18.00 3.23×10-3 23.60 5.07×10-3

Brief Description of Seismic Refraction Survey (NEA)

S.N. Location Seismic Line Length (m) 1 Headworks Site SLD-1 230 2 Headworks Site SLD-2 115 3 Headworks Site SLD-3 170 4 Headworks Site SLD-4 115 5 Headworks Site SLD-5 115 6 Headworks Site SLD-6 115 7 Headworks Site SLD-7 345 8 Headworks Site SLD-8 115 9 Headworks Site SLD-9 115 10 Headworks Site SLD-10 115 11 Underground Powerhouse Site SLP-1 230 12 Underground Powerhouse Site SLP-2 115 13 Underground Powerhouse Site SLP-3 115 14 Underground Powerhouse Site SLP-4 115 15 Underground Powerhouse Site SLP-5 115 16 Underground Powerhouse Site SLP-6 230 17 Underground Powerhouse Site SLP-7 115 18 Underground Powerhouse Site SLP-8 55 19 Underground Powerhouse Site SLP-9 55 20 Underground Powerhouse Site SLP-10 55 21 Underground Powerhouse Site SLP-11 55 22 Underground Powerhouse Site SLP-12 115 23 Underground Powerhouse Site SLP-13 55 24 Underground Powerhouse Site SLP-14 55

vii

Annex III

Rock Mass Rating System (After Bieniawski, 1989)

viii

Rock Mass Rating System (After Bieniawski, 1989)

A. CLASSIFICATION PARAMETER AND THEIR RATINGS Parameter Range of values

1 Strength of intact

rock

Point load strength index

>10 MPa 4-10 MPa 2-4 MPa 1-2 MPa For this low range uniaxial comp. test is

preffered Uniaxial comp.

strength >250 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 3 3 Spacing of discontinuities >2 m 0.6-2 m 200-600 mm 60-200 mm <60 mm

Rating 20 15 10 8 5 4 Condition of discontinuities

(see E) Very rough surface Not continuous No separation Unweathered wall

Slightly rough surface Separation <1mm Slightly weathered wall

Slightly rough surfaces Separation <1mm Highly weathered walls

Silckensided surface or Gouge <5mm thick or separation 1-5mm continuous

Soft gouge >5mm thick or

Separation >5mm continuous

Rating 30 25 20 10 0 5 Ground

water Inflow per 10 m tunnel length (l/m)

None < 10 10 – 25 25 – 125 > 125

(joint water press)/ (Major principal σ)

0 < 0.1 0.1 – 0.2 0.2 – 0.5 > 0.5

General condition Completely dry Damp Wet Dripping Flowing Rating 15 10 7 4 0

B. RATING ADJUSTMENT FOR DISCONTINUITY ORIENTATION (see F) Strike and dip orientation Very favourable Favourable Fair Unfavourable Very unfavourable

Rating Tunnels & mines 0 -2 -5 -10 -12 Foundation 0 -2 -7 -15 -25

Slopes 0 -5 -25 -50 C. ROCK MASS CLASS DETERMINED FROM TOTAL RATINGS Rating 100 - 81 80 - 61 60 - 41 40 - 21 < 21 Class number I II III IV V description Very good rock Good rock Fair rock Poor rock Very poor rock D. MEANING OF ROCK MASS CLASSES Class number I II III IV V Average stand-up time 20 yrs for 15m span 1 yr for 10m span 1week for 5m span 10 hrs for 2.5m span 30 min for 1m span Cohesion of rock mass (kPa) > 400 300 - 400 200 - 300 100 - 200 < 100 Friction angle of rock mass > 45º 35º - 45º 25º - 35º 15º - 25º < 15º E. GUIDELINES FOR CLASSIFICATION OF DISCONTINUITY CONDITIONS* Discontinuity length(persistency) Rating

< 1m 6

1 - 3 m 4

3 - 10 m 2

10 - 20 m 1

> 20 m 0

Separation (aperture) Rating

None 6

< 0.1 mm 5

0.1 - 1.0 mm 4

1 - 5mm 1

> 5mm 0

Roughness Rating

Very rough 6

Rough 5

Slightly rough 3

Smooth 1

Slickensided 0

Infilling (gouge) Rating

None 6

Hard filling <5mm 4

Hard filling >5mm 2

Soft filling <5mm 2

Soft filling >5mm 0

Weathering Rating

Unweathered 6

Slightly weathered 4

Moderately weathered

3

Highly weathered 1

Decomposed 0

F. EFFECT OF DISCONTINUITY STRIKE AND DIP ORIENTATION IN TUNNIRLLING** Strike perpendicular to tunnel axis Strike parallel to tunnel axis

Drive with dip – Dip 45º - 90º Drive with dip – Dip 20º - 45º Dip 45º - 90º Dip 20º - 45º Very favourable Favourable Very unfavourable Fair

Drive against dip – Dip 45º - 90º Drive against dip – Dip 20º - 45º Dip 0º - 20º - irrespective of strike Fair Unfavourable Fair

* some conditions are mutually exclusive. For example, if infilling is present, the roughness of the surface will be overshadowed by the influence of the gouge. In such case use A.4 directly

** Modified after Wickham et al (1972).

ix

Annex IV

Classification of indivisual parameters used in the Quality Index Q (After Barton et al 1974)

x

Classification of indivisual parameters used in the Quality Index Q (After Barton et al 1974)

DESCRIPTION VALUE NOTES 1. ROCK QUALITY DESIGNATION A. Very poor B. Poor C. Fair D. Good E. Excellent

RQD 0 – 25

25 – 50 50 – 75 75 – 90 90 – 100

1. Where RQD is reported or measured as ≤ 10 (including 0), a nominal value of 10 is used to evaluate Q. 2. RQD intervals of 5, i.e. 100, 95, 90 etc. are sufficiently accurate.

2. JOINT SET NUMBER A. Massive, no or few joints B. One joint set C. One joint set plus random D. Two joint sets E. Two joint sets plus random F. Three joint sets G. Three joint sets plus random H. Four or more joint sets, random, heavily jointed, sugar cube, etc. J. Crushed rock, earthlike.

Jn 0.5 – 1.0

2 3 4 6 9

12 15 20

1. For intersections use (3.0 ×Jn) 2. For portals use (2.0 × Jn)

3. JOINT ROUGHNESS NUMBER a. Rock wall contact b. Rock wall contact before 10 cm shear A. Discontinuous joints B. Rough and irregular, undulating C. smooth undulating D. Slickensided unulating E. Rough or irregular, planar F. Smooth, planar G. Slickensided, planar c. No rock wall contact when sheared H. Zones containgin clay minerals thick enough to prevent roc wall contact J. Sandy, gravely or crushed zone thick enouht to prevent rock wall contact

Jr

4 3 2

1.5 1.5 1.0 0.5

1.0 1.0

1. Add 1.0 if the mean spacing of the relevant joint set is greater than 3 m. 2. Jr =0.5 can be used for planar, slickensided joints having lineations, provided that the lineation are oriented for minimum strength.

4. JOINT ALTERATION NUMBER a. Rock wall contact A. Tightly healed, hard, non-softening, impermeable filling B. Unaltered joint walls, surface staining only C. Slightly alrered join walls non-softening mineral coating, sandy particles, clay-

free disintegrated rock etc. D. Silty or sandy-clay coating, small clay-fraction (non-softening) E. Softening or low friction clay mineral coating, i.e. kaolinite, mica. Also chorite,

talc, gypsum and graphite etc, and small quantities of swelling clays. (Discontinuous coating, 1 – 2 mm or less)

b. Rock wall contact before 10 cm shear F. Sandy particles, clay-free, disintegrating rock etc. G. Strongly over-consolidated, non-softening clay mineral filling (continuous < 5

mm thick) H. Medium or low over-consolidation, softening clay mineral filling (continuous < 5

mm thick) I. Swelling clay filling, i.e. montmorillonite, (continuous < 5 mm thick). Values of Ja

depend on percent of swelling clay-size particles, and access to water. c. No rock wall contact when sheared. J. Zones or bands of disintegrated or crushed rock and clay (see G, H and I for clay

K. conditions) K. Zones or bands of silty- or clay, small clay fraction (non softening) K. Thick continuous zones or bands of clay (see G, H and I for clay conditions)

Ja

0.75 1.0 2.0

3.0 4.0

4.0 6.0

8.0

8.0 – 12.0

6.0, 8.0 or 8.0 – 12.0

0.5 10.0 -13.0

or 13.0 -20.0

Фr degrees (approx.) 25 - 35 25 - 30 20 - 25 8 - 16 25 - 30 16 - 24 12 - 16 6 - 12 6 - 24 6 – 24 1. Values of Фr, the residual friction angle are intended as an approximate guide to the mineralogical properties of the

xi

alteration products, if present. 5. JOINT WATER REDUCTION A. Dry excavation or minor inflow i.e. < 5 l/m locally B. Medium inflow or pressure, occasional outwash of joint fillings C. Large inflow or high pressure in competent rock with unfilled joints. D. Large inflow or high pressure E. Exceptionally high inflow or pressure at blasting, decaying with time F. Exceptionally high inflow or pressure

Jw 1.0

0.66 0.5

0.33 0.2-0.1

0.1-0.05

Approx. water pressure (kgf?cm2) <1.0 1.0-2.5 2.5-10.0 2.5-10.0 >10 >10 1. Factors C to F are crude estimates: increase Jw if drainage installed. 2. Special problems caused by ice formation are not considered.

6. STRESS REDUCTION FACTOR a. Weakness zones intersection excavation, which may cause loosening of rock mass

when tunnel is excavated A. Multiple occurrences of weakness zones containing clay or chemically

disintegrated rock, very loose surrounding rock any depth B. Single weakness zone containing clay, or chemically disintegrated rock

(excavation depth < 50 m) C. Single weakness zone containing clay, or chemically disintegrated rock

(excavation depth > 50 m) D. Multiple shear zones in competent rock (clay free), loose surrounding rock (any

depth) E. Single shear zone in competent rock (clay free), (depth of excavation < 50 m) F. Single shear zone in competent rock (clay free), (depth of excavation >50m) G. Loose open joints, heavily jointed or ‘sugar cube’, (any depth) b. Competent rock, rock stress problems H. Low stress, near surface I. Medium stress J. High stress, very tight structure (usually favorable to stability, may be unfavorable

to wall stability K. Mild rock burst (massive rock) L. Heavy rock burst (massive rock) c. Squeezing rock, plastic flow of incompetent rock crown below surface is less than

span width, under influence of high rock pressure. M. Mild squeezing rock pressure N. Heavy squeezing rock pressure d. Swelling rock, chemical swelling activity depending on presence of water O. Mild swelling rock pressure P. Heavy swelling rock pressure

SRF

10.0

5.0

2.5

7.5

5.0 2.5 5.0

2.5 1.0

0.5-2

5-10 10-20

5-10 10-20

5-10

10-15

1. Reduce these values of SRF by 25-50% but only if the relevant shear zones influence do not intersect the excavation. 2. For strongly anisotropic virgin stress field (if measured); when 5≤σ1/σ3≤10, reduce σc to 0.8σc and σt to 0.8σt. When σ1/σ3>10, reduce σc and σt to 0.6 σc and 0.6 σt , where σc= unconfined compressive strength, and σt = tensile strength (point load) and σ1 and σ3 are the major and minor principal stresses. σc/σ1 σt/σ1 >200 >13 200-10 13-0.66 10-5 0.66-0.33 5-2.5 0.33-0.16 <2.5 <0.16 3. Few cases records available where depth of crown suggest SRF increased from 2.5 to 5 for such cases (see H).

ADDITIONAL NOTES ON THE USE OF THESES TABLES When making estimates of the rock mass Quality (Q), the following guidelines should be followed in addition to the notes listed in the tables: 1. When borehole core is unavailable, RQD can be estimated from the number of joints per unit volume, in which the number of joints per meter for each joint set are added. A simple relationship can be used to convert this number to RQD for the case of clay free rock masses: RQD = 115-3.3 Jv (approx.), where Jv = total number of joints per m3 (0 < RQD < 100 for 35 > Jv > 4.5). 2. The parameter Jn representing the number of joint sets will often be affected by foliation, schistosity, slaty cleavage or bedding etc. If strongly developed, these parallel “joints” should obviously be counted as a complete joint set. However, if there are few ‘joints’ visible, or if only occasional breaks in the core are due to these feature, then it will be more appropriate to count them as ‘random’ joints when evaluating Jn 3. The parameters Jr and Ja (representing shear strength) should be relevant to the weakest significant joint set or clay filled discontinuity in the given zone. However, if the joint set or discontinuity with the minimum value of Jr / Ja is favourably oriented for stabilty, then a second, less favourably oriented joint set or discontinuity may sometimes be more significant, and its higher value of Jr/Ja should be used when evaluating Q. The value of Jr/Ja should in fact relate to the surface most likely to allow failure to initiate. 4. When a rock mass contains clay, the factor stress reduction factor (SRF) appropriate to loosening loads should be evaluated. In such cases the strength of the intact rock is of little interest. However, when jointing is minimal and clay is completely absent, the strength of the intact rock may become the weakest link, and the stability will then depend on the ratio rock-stress/rock-strength. A strongly anisotropic stress field is unfavourable for stability and is roughly accounted for as in note 2 in the table for SRF evaluation. 5. The compressive and tensile strength (σc and σt) of the intact rock should be evaluated in the saturated condition if this is appropriate to the present and future in situ condition. A very conservative estimate of the strength should be made for those rocks that deteriorate when exposed to moist or saturated conditions.

xiii

Annex V Mohr-Coulomb failure criteria

\

xiv

Mohr-Coulomb failure criteria for Ch 0+000 m to 0+100 m

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Mohr-Coulomb failure criteria for Surge Shaft

xxii

Mohr-Coulomb failure criteria for Inclined Shaft

xxiii

Mohr-Coulomb failure criteria for Powerhouse

xxiv

Mohr-Coulomb failure criteria for Tailrace tunnel

xxv

Annex VI

Rock Mass Rating (RMR) and Tunnel Quality Index (Q) along Headrace Tunnel

xxviii

Engineering Geological and Geotechnical Study of Upper Trisuli - 3a Hydroelectric Project, Nuwakot...

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