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
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
i
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
ii
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
External Examiner
Acting Head
Mr. Churna Bahadur Oli
Central Department of Geology
Senior Engineering Geologist Department of Electricity Development
…………………………………….
…………………………………….
Supervisor
Internal Supervisor
Mr. Jayandra Man Tamrakar
Dr. Kamala Kant Acharya
Manager (Senior Geologist)
Lecturer
Nepal Electricity Authority
Central Department of Geology iii
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
iv
“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.
v
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 vi
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
vii
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 viii
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! defined. ix
Bookmark
not
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 Figure 5.12: Stereographic projection of discontinuities measured along road cut section at down hill of Danda guan. ........................................................................... 47 Figure 5.13: Geological cross section from surge tank to tailrace tunnel. ................Error! Bookmark not defined. Figure 5.14: Detailed engineering geological map of powerhouse area. ..................Error! Bookmark not defined. Figure 5.15: Photograph of powerhouse site near Paire guan. ......................................... 53 Figure 5.16: Stereographic projection of discontinuities measured along road cut section 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 6.4: Support applied to stabilize Wedge No 7 ...................................................... 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 6.8: Support applied to stabilize Wedge No 7 ...................................................... 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 x
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 6.15: Support applied for Wedge No 6 ................................................................. 75 Figure 6.16: Support applied for Wedge No 8 ................................................................. 75 Figure 6.17: Stereoplot of major joint set in powerhouse cavern .................................... 76 Figure 6.18: Wedge expected in powerhouse cavern. ...................................................... 76 Figure 6.19: Support applied for Wedge No 6 ................................................................. 77 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
xi
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 xii
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
xiii
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 xiv
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,
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.
2
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.
3
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) 4
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 v.
El. 726 m Power Generation
Minimum power generation
43.75 MW
Annual average energy
489.76 GWh (gross) 5
6
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
7
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.
8
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
9
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 SubHimalaya, 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 10
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 11
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. 12
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.
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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. ¾ 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. ¾ 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. ¾ 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. ¾ 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. ¾ 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 PaleozoicMesozoic 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 14
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. ¾ 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. ¾ 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. ¾ 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. ¾ 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.
15
¾ 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. ¾ 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. ¾ 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. ¾ 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. ¾ 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. 16
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 GorkhaNawakot 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.
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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 lowgrademetamorphism, 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).
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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.
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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
Phulchauki Group
Formation. Godavari Limestone Chitlang Formation Chandragiri Limestone Sopyang Formation Tistung Formation
Kathmandu Complex
Thickness (m) 300 1000
Main Lithology Limestone, dolomite Slate
2000
Limestone
200
Slate, calc-phyllite
3000
Metasandstone, Phyllite
----------------------------Transition------------------------------------------Markhu Formation 1000 Marble, schist Kulekhani Formation 2000 Quartzite, schist Bhimphedi Group
Chisapani Quartize
400
White quartzite
Kalitar Formation Bhainsedobhan Marble
2000
Schist, quartzite
800
Marble
Raduwa Formation
1000
Garnet-schist, quartz
---------------------------------------Mahabharat Thrust--------------------------------Upper Nawakot Group
Robang Formation
200-1000
Phyllite, quartzite
Malekhu Formation
800
Limestone, dolomite
Benighat Slate
500-3000
Slate, argillite, dolomites
--------------------------Erosional Unconformity (?)------------------------Nawakot Complex 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
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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 SESW 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.
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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.
22
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.
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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
25
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) 26
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 27
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)
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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
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)
31
(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. 32
34
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
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 37
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.
38
Figure 5.5: Photograph of Intake Portal facing 250º
Figure 5.6: Stereographic projection of discontinuities measured at Intake Portal 39
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 40
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: 117º142º/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.
41
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 The tunnel alignment from Ch 0+707 m to Ch 0+876 m runs with an azimuth 30º.The 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.
Ch 0+876 m to Ch 1+938 m The tunnel alignment from Ch 0+876 m to Ch 1+938 m runs with an azimuth 30º.The 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 The tunnel alignment from Ch 1+938 m to Ch 2+600 m runs with an azimuth 48º.The rock mass is milky white, medium- to thick-banded, slightly weathered, strong augen 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 45
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.
Figure 5.11: Stereographic projection of discontinuities measured along road cut section downhill of Diyale guan. Ch 2+600 m to Ch 4+142 m The tunnel alignment from Ch 2+600 m to Ch 4+142 m runs with an azimuth 15º.The rock mass is milky white, medium- to thick-banded, slightly weathered, strong augen 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.
46
Figure 5.12: Stereographic projection of discontinuities measured along road cut section 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 47
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
1+983m to 2+600m
662m
Gneiss
61-68
II
Good
2+600m to 4+142m
1542m
Gneiss
61-68
II
Good
Table 5.2: Rock mass classification along headrace tunnel based on Q value. Chainage
Length
Rock type
Q
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
1+983m to 2+600m
662m
Gneiss
9.90
III
Fair
2+600m to 4+142m
1542m
Gneiss
7.48
III
Fair
48
Rock Class
Rock catagory
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.
Figure 5.16: Stereographic projection of discontinuities measured along road cut section 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.
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 sandysilty 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.
54
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. ¾ 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. ¾ 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. ¾ Elastic and Plastic parameter are taken form empirical relation proposed by Hoek et al. (1995)
55
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
56
Table 6.1: Estimation of in situ deformation of rock along underground structures Structure
Headrace tunnel
Chainage 0+000m 0+100 m 0+100m – 0+200m 0+200m – 0+707m 0+707m – 0+876m 0+876m 1+938m 1+938m – 2+600m 2+600m – 4+142m
Surge shaft Inclined shaft Power house Tailrace tunnel
Em= 2RMR100 (GPa)
Em= 10(RMR(GPa)
Em= 25log10Q (GPa)
Average Em (GPa)
5.20
11.88
17.90
14.89
35
3.75
4.23
14.35
9.28
Gneiss
66
13.69
28.41
30.20
Gneiss
35
3.75
4.23
14.35
9.28
Gneiss
53
4.58
11.89
16.52
14.20
Gneiss
65
9.90
30
24.89
27.44
Gneiss
63
7.48
26
21.85
23.92
Gneiss
63
5.83
26
19.14
22.57
Gneiss
63
5.83
26
19.14
22.57
Gneiss
65
7.90
30
22.44
26.22
Gneiss
65
7.90
30
22.44
26.22
Rock type
RMR
Schist
53
Schist
Q
10)/40
32
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 57
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, 58
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
Headrace tunnel
0+000m 0+100 m 0+100m – 0+200m 0+200m – 0+707m 0+707m – 0+876m 0+876m 1+938m 1+938m – 2+600m 2+600m – 4+142m
Surge shaft Inclined shaft Power house Tailrace tunnel
Rock type
z (m)
γ (MN/m3)
Em (GPa)
σv (MPa)
k
σh (MPa)
Schist
75
0.027
14.89
2.025
1.744
3.531
Schist
142.5
0.027
9.28
3.847
0.771
2.966
Gneiss
225
0.027
30.20
6.075
1.401
8.511
Gneiss
202.5
0.027
9.28
5.467
0.636
3.476
Gneiss
210
0.027
14.20
5.670
0.823
4.665
Gneiss
165
0.027
27.44
4.455
1.606
7.156
Gneiss
300
0.027
23.92
8.100
0.976
7.902
Gneiss
75.75
0.027
22.57
2.045
2.493
5.100
Gneiss
277.5
0.027
22.57
7.492
0.977
7.323
Gneiss
202.5
0.027
26.22
5.467
1.339
7.326
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. 59
Table 6.3: Damage index of rock mass along underground structures. Structure
Chainage
Headrace tunnel
0+000m 0+100 m 0+100m – 0+200m 0+200m – 0+707m 0+707m – 0+876m 0+876m 1+938m 1+938m – 2+600m 2+600m – 4+142m
Surge shaft Inclined shaft Power house Tailrace tunnel
Rock type
σv (MPa)
k
σc (MPa)
σmax (MPa)
Di
Schist
2.025
1.744
35
8.569
0.245
Schist
3.847
0.771
35
5.0511
0.144
Gneiss
6.075
1.401
100
19.4582
0.195
Gneiss
5.467
0.636
100
4.9640
0.049
Gneiss
5.670
0.823
100
8.3292
0.083
Gneiss
4.455
1.606
100
17.0091
0.170
Gneiss
8.100
0.976
100
15.6168
0.156
Gneiss
2.045
2.493
100
13.2495
0.132
Gneiss
7.492
0.977
100
14.4670
0.145
Gneiss
5.467
1.339
100
16.4939
0.165
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 60
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 ½
σt = ½ σc (mb – (mb2 + 4s) ) ……………………………….(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, 61
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 Structure
Chainage
Headrace tunnel
0+000m 0+100 m 0+100m 0+200m 0+200m 0+707m 0+707m 0+876m 0+876m 1+938m 1+938m 2+600m 2+600m 4+142m
Surge shaft Inclined shaft Power house Tailrace tunnel
Rock type
Maximum rock cover, z (m)
GSI
σc (MPa)
mi
mb
s
Schist
75
48
35
12
1.8734
0.0031
Schist
142.5
30
35
12
0.9850
0.0042
Gneiss
225
61
100
28
6.9542
0.1312
Gneiss
202.5
30
100
28
2.2984
0.0042
Gneiss
210
48
100
28
4.3713
0.0031
Gneiss
165
60
100
28
6.7102
0.0117
Gneiss
300
58
100
28
6.2476
0.0094
Gneiss
75.75
58
100
28
6.2476
0.0094
Gneiss
277.5
58
100
28
6.2476
0.0094
Gneiss
202.5
60
100
28
6.7102
0.0117
Gneiss
150
60
100
28
6.7102
0.0117
62
Table 6.5: Values of mi for intact rock (Marrinos and Hoek, 2001)
63
Table 6.6: Analysis of rock strength using Roclab. Structure
Hoek-Brown classification
0
Failure envelope range
MohrCoulomb fit
Rock mass parameters
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
Intact uniaxial compressive strength (σc, MPa) GSI mi Disturbance factor (D)
35
35
100
100
100
100
100
100
100
100
100
48 12 0
30 12 0
61 28 0
30 28 0
48 28 0
60 28 0
58 28 0
58 28 0
58 28 0
60 28 0
60 28 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 a
0.0031 0.507
0.0004 0.522
0.0131 0.503
0.0004 0.522
0.0031 0.507
0.0117 0.503
0.0094 0.503
0.0094 0.503
0.0094 0.503
0.0117 0.503
0.0117 0.503
Application
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
Tunnel
σ3max (MPa) Unit weight (MN/m3) Overburden (m)
1.0195 0.027 75
1.8182 0.0027 142.5
3.1762 0.027 225
2.7663 0.027 202.5
2.9304 0.027 210
2.3701 0.027 165
4.0028 0.027 300
1.1373 0.027 75.75
3.8542 0.027 277.5
2.8732 0.027 202.5
2.1670 0.027 150
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
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
1.875
0.602
11.325
1.720
5.356
10.701
9.550
9.550
9.550
10.701
10.701
6.375
4.215
35.857
18.678
27.611
35.135
33.744
33.744
33.744
35.135
35.135
5272.71
1870.83
18836.49
3162.28
8912.51
17782.79
15848.93
15848.93
15848.93
17782.79
17782.79
Uniaxial compressive strength (MPa) Global strength (MPa) Deformation modulus (MPa)
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
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).
66
Figure 6.3: Support applied to stabilize Wedge No 6
Figure 6.4: Support applied to stabilize Wedge No 7 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).
67
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
68
Figure 6.7: Support applied to stabilize Wedge No 6 and 8
Figure 6.8: Support applied to stabilize Wedge No 7 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 69
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 Ch 2+600 m is plotted in streonet (Figure 6.9). The major joints and tunnel alignment 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.
70
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 71
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). 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 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
72
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 Ch 4+142 m is plotted in streonet (Figure 6.13). The major joints and tunnel alignment 73
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). 74
Figure 6.15: Support applied for Wedge No 6
Figure 6.16: Support applied for Wedge No 8 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 75
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 76
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).
Figure 6.19: Support applied for 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 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).
77
Figure 6.20: Support applied for Wedge No 7 and 8
78
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. 79
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
Excellent1 (RQD>90)
Steel sets
Rockbolts3
shotcrete
Boring machine
None to occasional light set, rock load (0.0-0.2)B
None to occasional
None to occasional local application
None to occasional light set, rock load (0.0-0.3)B Occassional light sets to pattern on 5- to 6-ft center. Rock load (0.0 to 0.4)B Light sets 5- to 6- ft center. Rock load (0.3 to 0.6)B Light to medium sets, 5- to 6–ft center. Rock load (0.4 to 1.0)B Light to medium sets, 4- to 5-ft center. Rock load (0.6 to 1.3)B Medium circular sets on 3- to 4-ft center. Rock load (1.0 to 1.6)B Medium to heavy circular sets on 2to 4-ft center. Rock load (1.3 to 2.0)B
None to occasional Occasional to pattern on 5- to 6-ft center. Pattern, 5- to 6ft centers Pattern, 4- to 6ft center Pattern, 3- to 5ft center Pattern, 3- to 5ft center
None to occasional local application 2 to 3 in
4-in. or more crown and sides. 4- to 6-in. on crown and sides. Combine with bolts.
Pattern, 2- to 4ft center
6-in. or more on crown and sides. Combine with bolts
Boring machine
Medium to heavy circular sets on 2ft center. Rock load (1.6 to 2.2)B
Pattern, 2- to 3ft center
Conventional
Heavy circular sets on 2-ft center. Rock load (1.6 to 2.2B)
Pattern, center
Boring machine
Very heavy circular sets on 2-ft center. Rock load up to 250 ft.
Pattern, 2- to 3ft center
Conventional
Very heavy circular sets on 2-ft center. Rock load up to 250 ft.
Pattern, 2- to 3ft center
Conventional Good1 (75
Boring machine Conventional
Fair (50
2
Poor (25
Very poor3 (RQD<25, excluding squeezing or swelling ground)
Very poor3 (RQD<25, squeezing or swelling ground)
Alternative support systems
Tunneling method
Boring machine Conventional Boring machine Conventional
3
3-ft
None to occasional local application 2 to 3 in. Occasional local application 2 to 3 in. 2- to 4-in. crown
6-in. or more on whole section. Combine with medium sets. 6-in. or more on whole section. Combine with medium sets. 6-in. or more on whole section. Combine with heavy sets 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.
80
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. Rock type
RQD (%)
Rock mass quality
0+000m 0+100 m
Schist
64
Fair
0+100m 0+200m
Schist
45
Poor
0+200m 0+707m
Gneiss
80
Good
0+707m 0+876m
Gneiss
45
Poor
Gneiss
58
Fair
Gneiss
73
Fair
Gneiss
70
Fair
Surge shaft
Gneiss
50
Poor
Inclined shaft
Gneiss
50
Poor
Gneiss
55
Fair
Gneiss
55
Fair
Structure
Headrace tunnel
Chainage
0+876m 1+938m 1+938m 2+600m 2+600m 4+142m
Power house Tailrace tunnel
Proposed support system Steel set
Rock bolts
Shotcrete
Light to medium sets on 5 to 6ft center Medium to heavy circular sets on 2- to 4-ft center. Rock load (1.3 to 2.0)B Light sets 5- to 6- ft center. Rock load (0.3 to 0.6)B Medium to heavy circular sets on 2- to 4-ft center. Rock load (1.3 to 2.0)B Light to medium sets on 5 to 6ft center Light to medium sets on 5 to 6ft center Light to medium sets on 5 to 6ft center Medium to heavy circular sets on 2- to 4-ft center. Rock load (1.3 to 2.0)B Medium to heavy circular sets on 2- to 4-ft center. Rock load (1.3 to 2.0)B Light to medium sets on 5 to 6ft center Light to medium sets on 5 to 6ft center
Pattern, 5 to 6ft center
Occasional local application 2 to 3in. 6-in. or more on crown and sides. Combine with bolts Occasional local application 2 to 3 in. 6-in. or more on crown and sides. Combine with bolts Occasional local application 2 to 3in. Occasional local application 2 to 3in. Occasional local application 2 to 3in. 6-in. or more on crown and sides. Combine with bolts 6-in. or more on crown and sides. Combine with bolts Occasional local application 2 to 3in. Occasional local application 2 to 3in.
Pattern, 2- to 4ft center Pattern, 5- to 6ft centers Pattern, 2- to 4ft center Pattern, 5 to 6ft center Pattern, 5 to 6ft center Pattern, 5 to 6ft center Pattern, 2- to 4ft center Pattern, 2- to 4ft center Pattern, 5 to 6ft center Pattern, 5 to 6ft center
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.
81
Table 6.9: Geomechanics classification guide for excavation and support in rock tunnels after Bieniawski (1989) Support Rock Mass
Very Good Rock, I RMR: 81-100 Good Rock II RMR: 61-80 Fair Rock, III RMR: 41-60 Poor Rock, IV RMR: 21-40
Very Poor Rock V, RMR; <20
Excavation
Full Face 3m advance Full face 1.0-1.5 m advance. Complete support 20 m from face Top heading and bench 1.53m advance in heading. Commence support after each blast. Complete support 10m from face. Top heading and bench, 11.5m advance in heading. Install support concurrently with excavation 10m from face Multiple drifts. 0.5-1.5m advance in top heading. Install support concurrently with excavation. Shotcrete as soon as possible after blasting
Rockbolts (20mm dia. Fully bonded
Shotcrete
Steel sets
Generally no support required except for occasional spot bolting Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh
50 mm in crown where required
None
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
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
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 =
Excavation span, diameter or height (m) Excavation Support Ratio
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.
82
Table 6.10: Estimation of excavation and support in underground structures based on RMR. Structure
Chainage
0+000m 0+100 m
Rock type
Schist
R M R
Rock class
53
Fair (III)
Schist
35
Poor (IV)
0+200m – 0+707m
Gneiss
66
Good (II)
35
Poor (IV)
53
Fair (III)
Full face 1.0-1.5 m advance. Complete support 20 m from face
0+707m – 0+876m
0+876m 1+938m
Gneiss
Gneiss
Gneiss
65
2+600m – 4+142m
Gneiss
63
Good (II)
Full face 1.0-1.5 m advance. Complete support 20 m from face
Gneiss
63
Good (II)
Full face 1.0-1.5 m advance. Complete support 20 m from face
Gneiss
63
Good (II)
Full face 1.0-1.5 m advance. Complete support 20 m from face
Gneiss
65
Good (II)
Full face 1.0-1.5 m advance. Complete support 20 m from face
Gneiss
65
Good (II)
Full face 1.0-1.5 m advance. Complete support 20 m from face
Power house Tailrace tunnel
Top heading and bench, 11.5m advance in heading. Install support concurrently with excavation 10m from face Top heading and bench 1.53m advance in heading. Commence support after each blast. Complete support 10m from face.
Good (II)
Inclined shaft
Full face 1.0-1.5 m advance. Complete support 20 m from face
1+938m – 2+600m
Surge shaft
Top heading and bench 1.53m advance in heading. Commence support after each blast. Complete support 10m from face. Top heading and bench, 11.5m advance in heading. Install support concurrently with excavation 10m from face
0+100m – 0+200m
Headrace tunnel
Excavation
83
Support Rock bolts
Shotcrete
Steel ribs
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
100-150mm in crown, and 100mm in sides
Light ribs spaced 1.5m where required
50 mm in crown where required
None
100-150mm in crown, and 100mm in sides
Light ribs spaced 1.5m where required
50-100mm in crown, 30 mm in side wall
None
50 mm in crown where required
None
50 mm in crown where required
None
50 mm in crown where required
None
50 mm in crown where required
None
50 mm in crown where required
None
50 mm in crown where required
None
Systematic bolts 45m long, spaced 11.5m in crown and walls with wire mesh Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh Systematic bolts 45m long, spaced 11.5m in crown and walls with wire mesh Systematic bolts 4m long, spaced 1.5-2m in crown and walls with mesh in crown. Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh Locally bolts in crown, 3m long spaced 2.5 m with occasional mesh
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=
(2 + 0.15 B ) ESR
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 84
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 =
2 Jn × Q −1 / 3 3 Jr
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) 85
Table 6.12: Estimation of rock support based on Q Structure
Chainage
Rock type
Max. unsupporte d span (m)
Proof (Kg/cm2)
Q
Rock class
6.19
0.768
III
Support category
0+000m 0+100 m
Schist
5.20
Fair (III)
0+100m – 0+200m
Schist
3.75
Poor (IV)
5.43
0.856
IV
0+200m – 0+707m
Gneiss
13.69
Good (II)
9.11
0.556
II
0+707m – 0+876m
Gneiss
3.75
Poor (IV)
5.43
0.856
IV
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
Headrace tunnel
Support Rock bolts
Shotcrete
Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 1.5m (center to center) spacing Spot bolting, 2m long, 2.5m (center to center) spacing Systematic bolting, 2m long, 1.5m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing Systematic bolting, 2m long, 2m (center to center) spacing
100mm fibrereinforced shotcrete at crown and wall 150mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 150mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall 100mm fibrereinforced shotcrete at crown and wall
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 86
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 Greatest of: (a) 2 × bolt spacing (b) 3 × thickness of critical and potentially unstable rock blocks 1 (c) For element above the spring line:
Minimum length and maximum spacing
Minimum length
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 Least of:
Maximum spacing
(a) 0.5 × bolt length (b) 1.5 × width of critical and potentially unstable rock blocks 1 (c) 2.0m 2
Minimum 0.9m to 1.2m spacing
Notes: 1.
2.
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). Greater spacing than 2.0m makes attachment of surface support elements (e.g., weld-mesh or chain-link mesh) difficult.
87
Table 6.14: Estimation of rock support based on empirical design recommendation to U.S. corps of engineers Structure
Chainage
0+000m 0+100 m
0+100m – 0+200m
0+200m – 0+707m
Headrace tunnel
0+707m – 0+876m
0+876m 1+938m
1+938m – 2+600m
2+600m – 4+142m
Surge shaft
Rock type
Recommendation for rock bolt reinforcement Minimum length
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
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
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
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
Gneiss
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
Gneiss
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
Gneiss
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
Gneiss
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
88
Maximum spacing 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 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 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 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 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 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 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 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
Minimum spacing
0.8m to 1.25m
0.8m to 1.25m
0.8m to 1.25m
0.8m to 1.25m
0.8m to 1.25m
0.8m to 1.25m
0.8m to 1.25m
0.8m to 1.25m
Inclined shaft
Power house
Tailrace tunnel
Gneiss
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
Gneiss
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
Gneiss
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
(c) 2.0m 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 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 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
0.8m to 1.25m
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. A. 1. 2. 3. B. 4 C. 5. D. 6. E. 7.
Structure Headrace tunnel Poor Rocks (Rock Class – IV) Fair Rocks (Rock Class – III) Good Rocks (Rock Class – II) Surge Shaft Fair Rocks (Rock Class – IV) Inclined Shaft Fair Rocks (Rock Class – IV) Powerhouse Fair Rocks (Rock Class – IV) Tailrace Tunnel Fair Rocks (Rock Class – IV)
Fibre reinforced shotcrete Crown Sidewall
Rock bolt Length
spacing
150 mm
150mm
2.5 m
1.5 m × 1.5 m
100 mm
100 mm
2.5 m
2m×2m
2.5 m
2.5 m × 2.5 m
150 mm
150 mm
8m
2m×2m
100 mm
100 mm
2.5 m
2m×2m
150 mm
150 mm
8m
2m×2m
100 mm
100 mm
2.5 m
2m×2m
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.
89
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.
¾
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
¾
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.
¾
Gneiss unit consist of milky white, medium- to thick-banded augen gneiss with occassional partings of light grey to greenish grey schist.
¾ In general, attitude of foliation plane of rock with in project area varies from NNESSW to NNW-SSE dipping 10º-30º towards NE-SW.
¾ The Main Boundary Thrust (MBT) passes about 80 km south and Main Central Thrust (MCT) passes about 25 km north of the project area.
¾ The bedrock peak ground acceleration value of the project area is 250-300 gal and estimated design coefficient is 0.13-0.15.
¾ 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.
¾ 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.
90
¾ 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.
¾ 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.
¾ 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.
¾ 95% of the headrace tunnel will passes through augen gneiss and only 5% will passes through schist.
¾ 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.
¾ 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.
¾ The attitude of foliation is almost perpendicular to tunnel alignment dipping to NENW 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.
¾ 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.
¾ Average in situ deformation modulus (Em) along the underground structures ranges from 9.28 to 30.20 GPa.
91
¾ 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.
¾
Damage index (Di) along underground structure is less than 0.4 (0.049 – 0.195). Hence, the rock mass behaves as an elastic.
¾
Support design for various underground structures based on different systems suggests the combination of local to systematic bolting and reinforced shotcrete as per requirement.
92
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Marinos, P., Hoek, E., 2001. Estimating the geotechnical properties of heterogeneous rock masses such as Flysch. Bull. Engg. Geol. Env. 60, pp. 85-92. Medlicott, H.B., 1875. Note on the Geology of Nepal. Records of the Geological Survey of India 8, pp. 93-101. Munthe, J., Dongol, B., Hutchison, J.H., Keans, W.F., Munthe, K., West, R.M., 1983. New fossil discoveries from the Miocene of Nepal include a hominoid. Nature, 303, pp. 331-333. Nepal Electricity Authority, 2006. Detailed Project Report of Upper Trisuli-3A Hudroelectric Project. Unpublished report. Kathmandu, Nepal.. Pandey, M.R., Tandukar, R.P., Avouac, J.P., Lavé, J., Massot, J.P., 1995. Interseismic strain accumulation on the Himalayan crustal ramp (Nepal). Geophysical Research Letters, 22, pp. 751-754. Power Development of Nepal, Nepal Electricity Authority, fiscal year 2007/08 – A year in review. Rai, S.M., 1998. étude structurale, métamorphique, géochimique et radiochronologique des nappes de Katmandou et du Gosainkund, Himalaya de Népal central (Thés d'université thesis), Univ Joseph-Fourier, Grenoble, 244p. Sakai, H., 1983. Geology of the Tansen Group of the Lesser Himalaya in Nepal. Memoirs of the Faculty of Science Kyushu University Series D Geology, 25, pp. 27-74. Sakai, H., 1985. Geology of the Kali Gandaki Supergroup of the Lesser Himalayas in Nepal. Memoirs of the Faculty of Science Kyushu University Series D Geology, 25, pp. 337-397. Seeber, L., Armbruster, J.G., 1981. Great detachment earthquakes along the Himalayan arc and long-term forecasting. Earthquake prediction - an International review maurice Ewing Series 4. American Geophysical Union, pp. 259-277.
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Serafim, J.L, Pereira, J.P.,1983. Consideration of the geomechanical classification of Bieniawski. Proc. Int. Symp. on engineering geology and Underground Construction, Lisbon, 1(II), pp.33-44. Sharma, C.K., 1973. In: Geology of Nepal. Mani Ram Sharma, Kathmandu, 189p. Sheory, P.R., 1994. A theory for in situ stresses in isotropic and transversely isotropic rock. Int. J. Rock Mech. Min. Sci. & Geomech., Abstr., 31(1), pp. 23-34. Stöcklin, J., 1980. Geology of Nepal and its regional Frame. Journal of the Geological Society of London, 137, pp. 1-34. Stöcklin, J., Bhattarai, K.D., 1977. Geology of the Kathmandu area and central Mahabharat range, Nepal Himalaya. Report of Department of Mines and Geology/ UNDP (unpublished), 86p. Terzaghi, K., 1946. Rock defects and loads on tunnel supports. In rock tunneling with steel supports, (eds R. V. Proctor and T. C. White) 1, 17-99. Youngstown, OH: Commercial Shesring and Stamping Company. Pp. 17-99. West, R.M., Munthe, J., Lukacs, J.R., Shrestha, T.B., 1975. Fossil mollusca from the Siwaliks of Eastern Nepal. Current Science, 44, pp. 497-498. West, R.M., Lukacs, J.R., Munthe, J., Hussain, S.T., 1978. Vertebrate fauna from Neogene Siwalik group Dang valley, western Nepal. Journal of Paleontology, 52, pp. 1015-1022. West, R.M., Munthe, J., 1981. Neogene vertebrate paleontology and stratigraphy of Nepal. Journal of Nepal Geological Society, 1, pp. 1-14. Valdiya, K.S., 1995. Proterozoic sedimentation
and Pan-African
geodynamic
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.
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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
98
99
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 Annex VI Tunnel
Mohr-Coulomb failure criteria Rock Mass Rating (RMR) and Tunnel Quality Index (Q) along Headrace
Annex I Calculation of Uniaxial Confined Strength of Schist using Schimdt Hammer Value and Laboratory Test result of Core Samples of Gneiss.
ii
Calculation of Uniaxial Confined Strength of Schist using Schimdt Hammer Value. S.N. 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53
Schmidt Hammer Value 38 40 34 42 40 32 24 22 42 28 22 32 50 42 26 22 40 42 34 36 32 36 30 32 28 26 32 42 42 44 44 56 38 42 37 44 32 38 32 38 26 38 32 28 46 46 22 32 32 24 34 28 24
α
UCS (MPa)
S. N
+90 -90 -90 +90 -90 +90 +90 +90 -90 +90 +90 +90 +90 +90 +90 +90 +90 +90 -90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 -90 +90 -90 -90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90
80 55 40 100 55 58 38 34 55 45 34 58 150 100 42 34 90 100 40 75 58 75 52 50 45 42 58 100 100 110 110 208 80 100 77 110 58 80 35 80 25 48 58 45 125 125 34 58 58 38 65 45 44
89 90 91 92 93 94 95 96 97 98 99 100 101 102 103 104 105 106 107 108 109 110 111 112 113 114 115 116 117 118 119 120 121 122 123 124 125 126 127 128 129 130 131 132 133 134 135 136 137 138 139 140 141
Schmidt Hammer Value 24 18 22 42 58 18 32 24 38 36 29 28 43 52 44 38 44 48 58 50 48 54 46 40 52 54 44 54 38 18 46 30 50 20 24 40 42 22 22 30 46 48 26 42 46 39 47 54 32 48 39 48 44
iii
UCS (MPa)
Mean UCS (MPa)
α -90 +90 +90 +90 +90 +90 -90 -90 +90 +90 +90 +90 +90 -90 +90 +90 +90 +90 -90 -90 +90 +90 +90 +90 -90 -90 -90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 -90 +90 +90 -90 +90 +90 +90 +90 +90 +90
23 26 34 100 230 26 35 23 80 75 48 45 105 105 110 80 110 137 139 58 137 195 125 100 105 115 68 198 80 26 125 52 150 30 28 90 100 34 34 52 125 137 42 60 125 88 80 154 58 137 88 137 110
35.15
54 55 56 57 58 59 60 61 62 63 64 65 66 67 68 69 70 71 72 73 74 75 76 77 78 79 80 81 82 83 84 85 86 87 88
30 40 48 14 20 22 18 16 24 16 24 44 17 18 18 25 20 16 35 15 18 22 22 34 36 34 34 38 34 32 42 54 47 47 39
+90 +90 +90 +90 +90 +90 +90 +90 +90 +90 -90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90 +90
52 90 137 21 30 34 26 24 38 24 23 68 25 26 26 40 30 24 68 23 26 34 34 65 75 65 65 80 65 58 100 195 135 135 88
142 143 144 145 146 147 148 149 150 151 152 153 154 155 156 157 158 159 160 161 162 163 164 165 166 167 168 169 170 171 172 173 174 175 176
54 44 46 54 30 38 42 38 47 54 34 42 52 42 43 20 40 26 38 40 3 30 28 22 26 20 28 28 36 24 24 42 44 40 30
+90 +90 +90 +90 -90 -90 +90 -90 -90 +90 -90 -90 -90 -90 -90 +90 +90 +90 -90 -90 -90 +90 +90 +90 +90 +90 +90 +90 -90 +90 +90 -90 -90 +90 +90
175 110 125 195 32 48 100 48 80 115 40 60 105 60 65 30 90 42 48 55 38 52 45 34 42 30 45 45 44 38 38 60 68 90 52
Laboratory Test Result for Core Samples of Gneiss (NEA) Depth m 31.60-32.00 46.33-46.90 59.00-59.55 69.55-70.00 26.00-26.50 35.23-35.70 46.00-46.70 58.00-58.65
Location
Powerhouse
Surge shaft
Specific Gravity
Absorption %
Unit Weight gm/cm3
2.645 2.66 2.65 2.70 2.66 2.67 2.68 2.64
0.5 0.3 0.3 0.4 0.3 0.3 0.3 0.4
2.6 2.6 2.6 2.6 2.6 2.6 2.6 2.6
iv
Uniaxial Compressive Strength kg/cm2 380.1 930.8 504.3 1006.2 880.4 955.9 883.4 1257.1
Young Modulus Eav N/mm2 2941.4 10962.6 4845.9 6235.8 6695.0 6766.2 18866.3 16913.4
Annex II General Description of Drill Holes, Summary of Constant Head Permeability Test Results and Brief Description of Seismic Refraction Survey
v
General Description of Drill Holes (NEA) Drill Hole No.
Drilling Machine
DH-1 DH-2 DH-3 DP-3 DP-4
Tone UD-5 Tone UD-5 Tone UD-5 Acker ‘Ace’ Acker ‘Ace’
Inclination & Direction Vertical Vertical Vertical Vertical Vertical
Location
Length (m)
Weir Axis (L/B) Weir Axis (R/B) Desander (R/B) Underground Powerhouse Surge Shaft
30.20 35.00 25.20 70.00 60.20
Summary of Constant Head Permeability Test Results (NEA) Drill Hole No.
Test Depth (m) 17.05 24.30 27.75 12.20 18.60 25.00 32.00 5.20 10.20 18.00 23.60
Location
DH-1
Weir Axis (L/B)
DH-2
Weir Axis (R/B)
DH-3
Desander
Permeability Value cm/s 1.01×10-1 6.80×10-3 5.00×10-3 8.20×10-2 6.82×10-2 3.66×10-3 1.58×10-3 2.22×10-2 4.62×10-2 3.23×10-3 5.07×10-3
Brief Description of Seismic Refraction Survey (NEA) S.N. 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24
Location Headworks Site Headworks Site Headworks Site Headworks Site Headworks Site Headworks Site Headworks Site Headworks Site Headworks Site Headworks Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site Underground Powerhouse Site
Seismic Line SLD-1 SLD-2 SLD-3 SLD-4 SLD-5 SLD-6 SLD-7 SLD-8 SLD-9 SLD-10 SLP-1 SLP-2 SLP-3 SLP-4 SLP-5 SLP-6 SLP-7 SLP-8 SLP-9 SLP-10 SLP-11 SLP-12 SLP-13 SLP-14
vi
Length (m) 230 115 170 115 115 115 345 115 115 115 230 115 115 115 115 230 115 55 55 55 55 115 55 55
Annex III Rock Mass Rating System (After Bieniawski, 1989)
vii
Rock Mass Rating System (After Bieniawski, 1989) A. CLASSIFICATION PARAMETER AND THEIR RATINGS Parameter Range of values 1 Strength Point load >10 MPa 4-10 MPa 2-4 MPa of intact strength index rock Uniaxial comp. >250 MPa 100-250 MPa 50-100 MPa strength Rating 15 12 7 2 Drill core quality RQD 90%-100% 75%-90% 50%-75% Rating 20 17 13 3 Spacing of discontinuities >2 m 0.6-2 m 200-600 mm Rating 20 15 10 4 Condition of discontinuities Very rough surface Slightly rough Slightly rough (see E) Not continuous No surface surfaces separation Separation <1mm Separation <1mm Unweathered wall Slightly weathered Highly weathered wall walls Rating 30 25 20 5 Ground Inflow per 10 m None < 10 10 – 25 water tunnel length (l/m) (joint water press)/ 0 < 0.1 0.1 – 0.2 (Major principal σ) General condition Completely dry Damp Wet Rating 15 10 7 B. RATING ADJUSTMENT FOR DISCONTINUITY ORIENTATION (see F) Strike and dip orientation Very favourable Favourable Fair Rating Tunnels & mines 0 -2 -5 Foundation 0 -2 -7 Slopes 0 -5 -25 C. ROCK MASS CLASS DETERMINED FROM TOTAL RATINGS Rating 100 - 81 80 - 61 60 - 41 Class number I II III description Very good rock Good rock Fair rock D. MEANING OF ROCK MASS CLASSES Class number I II III Average stand-up time 20 yrs for 15m span 1 yr for 10m span 1week for 5m span Cohesion of rock mass (kPa) > 400 300 - 400 200 - 300 Friction angle of rock mass > 45º 35º - 45º 25º - 35º E. GUIDELINES FOR CLASSIFICATION OF DISCONTINUITY CONDITIONS* Discontinuity length(persistency) < 1m 1-3m 3 - 10 m Rating 6 4 2 Separation (aperture) None < 0.1 mm 0.1 - 1.0 mm Rating 6 5 4 Roughness Rating Infilling (gouge) Rating Weathering Rating
Very rough 6 None 6 Unweathered 6
1-2 MPa 25-50 MPa 4 25%-50% 8 60-200 mm 8 Silckensided surface or Gouge <5mm thick or separation 1-5mm continuous
For this low range uniaxial comp. test is preffered 5-25 1-5 <1 Mpa Mpa MPa 2 1 0 <25% 3 <60 mm 5 Soft gouge >5mm thick or Separation >5mm continuous
10 25 – 125
0 > 125
0.2 – 0.5
> 0.5
Dripping 4
Flowing 0
Unfavourable -10 -15 -50
Very unfavourable -12 -25
40 - 21 IV Poor rock
< 21 V Very poor rock
IV 10 hrs for 2.5m span 100 - 200 15º - 25º
V 30 min for 1m span < 100 < 15º
10 - 20 m 1 1 - 5mm 1
Rough 5 Hard filling <5mm 4 Slightly weathered 4
> 20 m 0 > 5mm 0
Slightly rough Smooth Slickensided 3 1 0 Hard filling >5mm Soft filling <5mm Soft filling >5mm 2 2 0 Moderately Highly weathered Decomposed weathered 1 0 3 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).
viii
Annex IV Classification of indivisual parameters used in the Quality Index Q (After Barton et al 1974)
ix
Classification of indivisual parameters used in the Quality Index Q (After Barton et al 1974) DESCRIPTION 1. ROCK QUALITY DESIGNATION A. Very poor B. Poor C. Fair D. Good E. Excellent 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. 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 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, clayfree 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)
VALUE RQD 0 – 25 25 – 50 50 – 75 75 – 90 90 – 100 Jn 0.5 – 1.0 2 3 4 6 9 12 15 20 Jr 4 3 2 1.5 1.5 1.0 0.5
NOTES 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.
1. For intersections use (3.0 ×Jn) 2. For portals use (2.0 × Jn)
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.
1.0 1.0 Ja
Фr degrees (approx.)
0.75 1.0 2.0
25 - 35 25 - 30
3.0 4.0
20 - 25 8 - 16
4.0 6.0
25 - 30 16 - 24
8.0
12 - 16
8.0 – 12.0
6 - 12
6.0, 8.0 or 8.0 – 12.0 0.5 10.0 -13.0 or 13.0 -20.0
6 - 24 6 – 24 1. Values of Фr, the residual friction angle are intended as an approximate guide to the mineralogical properties of the
x
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
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
Jw 1.0 0.66 0.5 0.33 0.2-0.1 0.1-0.05
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
alteration products, if present. 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. 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 >200 200-10 10-5
σt/σ1 >13 13-0.66 0.66-0.33
5-2.5 <2.5
0.33-0.16 <0.16
3. Few cases records available where depth of crown suggest SRF increased from 2.5 to 5 for such cases (see H).
5-10 10-15
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.
xi
xii
Annex V Mohr-Coulomb failure criteria
\ xiii
Mohr-Coulomb failure criteria for Ch 0+000 m to 0+100 m
xiv
Mohr-Coulomb failure criteria for Ch 0+100 m to 0+200 m
xv
Mohr-Coulomb failure criteria for Ch 0+200 m to 0+707 m
xvi
Mohr-Coulomb failure criteria for Ch 0+707 m to 0+876 m
xvii
Mohr-Coulomb failure criteria for Ch 0+876 m to 0+938 m
xviii
Mohr-Coulomb failure criteria for Ch 0+938 m to 2+600 m
xix
Mohr-Coulomb failure criteria for Ch 2+600 m to 4+142 m
xx
Mohr-Coulomb failure criteria for Surge Shaft
xxi
Mohr-Coulomb failure criteria for Inclined Shaft
xxii
Mohr-Coulomb failure criteria for Powerhouse
xxiii
Mohr-Coulomb failure criteria for Tailrace tunnel
xxiv
Annex VI Rock Mass Rating (RMR) and Tunnel Quality Index (Q) along Headrace Tunnel
xxv
xxvii
xxviii
xxix
xxx
xxxi
xxxii
