Thesis 01

TABLE OF CONTENTS 1 CHAPTER 1 INTRODUCTION..........................................................1

1.1 GENERAL...................................................................................................................................... 1 1.2 PROJECT BACKGROUND........................................................................................................... 2 1.3 OBJECTIVES................................................................................................................................ 2

2 CHAPTER 2 GEOLOGY AND GEOMORPHOLOGY...................3 2.1 REGIONAL GEOLOGY................................................................................................................ 3 2.2 HEAD RACE TUNNEL ................................................................................................................. 3

3 CHAPTER 3 DESIGN OF THE TUNNEL SYSTEM.......................9 3.1 GENERAL...................................................................................................................................... 9 3.2 NORWEGIAN DESIGN APPROACH .......................................................................................... 9

3.2.1 Location of the Project............................................................................................................ 9 3.2.2 Orientation of Tunnel Axis.................................................................................................... 10 3.2.3 Shape of the Tunnels ............................................................................................................. 11 3.2.4 Dimensions ........................................................................................................................... 11

3.3 ROCK MASS CLASSIFICATION ALONG THE TUNNEL SYSTEM ...................................... 11 3.3.4 Q- Method............................................................................................................................. 11 3.3.2 RMR Method......................................................................................................................... 12

3.4 ROCK MASS RATING ALONG THE ORIGINAL ALIGNMENT ............................................ 12 3.5 DESIGN PRINCIPLES FOR UNLINED HEADRACE TUNNEL / PRESSURE SHAFTS DESIGN: .................................................................................................................................................. 13

3.5.1 Geological Restrictions: ....................................................................................................... 13 3.5.2 Topographical Restrictions: ................................................................................................. 13

3.6 OPTIMISATION OF WATERWAYS FOR UNLINED TUNNELS............................................ 14 3.7 SELECTION OF AN OPTIMUM LAYOUT ............................................................................... 15

3.7.1 Option 1:................................................................................................................................ 15 3.7.2 Option 2 ................................................................................................................................ 16 3.7.3 Option 3:............................................................................................................................... 16 3.7.4 Option 4:............................................................................................................................... 17 3.7.5 Selection of Best Option......................................................................................................... 17

4 CHAPTER 4.............. DESIGN OF AN UNDERGROUND POWER HOUSE .................................................................................................18

4.1 GENERAL.................................................................................................................................... 18 4.2 NORWEGIAN DESIGN APPROACH ........................................................................................ 18

4.2.1 Location ................................................................................................................................ 18 4.2.2 Orientation of Length axis:................................................................................................... 19 4.2.3 Shape..................................................................................................................................... 21 4.2.4 Dimensions ........................................................................................................................... 22

4.3 OPTIMISING OF CAVERN INTERNAL SPACE ...................................................................... 22 4.4 SPAN OF THE CAVERN................................................................................................................... 23

Transformer Layouts: ........................................................................................................................... 23 Escape Routes....................................................................................................................................... 24 Shafts .................................................................................................................................................... 24 Workshops and Repair facilities ........................................................................................................... 24 Control rooms and Social rooms .......................................................................................................... 25 Ventilation ............................................................................................................................................ 25

4.5 DIMENSIONING OF THE MACHINE HALL............................................................................ 25 4.5.1 Height ................................................................................................................................... 25 4.5.2 Span ...................................................................................................................................... 26

4.5.3 Erection Space ...................................................................................................................... 26 4.6 DIFFERENTPOWERHOUSE LAYOUT DESIGNS................................................................... 27

4.6.1 OPTION 1:............................................................................................................................ 27 4.6.2 OPTION 2:............................................................................................................................ 27 4.6.3 OPTION 3:............................................................................................................................ 27 4.6.4 OPTION 4:............................................................................................................................ 27

5 CHAPTER 5 DESIGN OF ROCK SUPPORTS...............................28 5.1 GENERAL APPROACH TO THE DESIGN OF ROCK SUPPORTS.......................................... 28 5.2 ROCK SUPPORT METHODS..................................................................................................... 28

5.2.1 Rock Bolting.......................................................................................................................... 28 5.2.2 Shotcreting............................................................................................................................ 29 5.2.3 Concrete Lining .................................................................................................................... 29 5.2.4 Steel Ribs .............................................................................................................................. 29 5.2.5 Grouting................................................................................................................................ 29

5.3 EVALUATION OF ROCK SUPPORTS ...................................................................................... 30 5.4 SUPPORT DESIGN FOR TUNNEL SYSTEM USING Q- METHOD .............................................. 31 5.5 SUPPORT DESIGN FOR PRESSURE SHAFT USING Q- METHOD.............................................. 32 5.3 SUPPORT DESIGN FOR POWER HOUSE CAVERN USING Q- METHOD ........................... 32

6 CHAPTER 6........................... NUMERICAL METHODS IN ROCK ENGINEERING ..................................................................................34

6.1 GENERAL.......................................................................................................................................... 34 6.2.1 Finite Element Method (FEM)..................................................................................................... 35 6.2.2 Finite Difference Method (FDM)................................................................................................ 35 6.2.3 Boundary Element Method (BEM) ............................................................................................. 35 6.2.4 Discrete Element Method (DEM) ................................................................................................ 36 6.2.5 Beam Element Method with Elastic support ................................................................................ 36 6.2.6 Hybrid Methods ........................................................................................................................... 36

6.3 SELECTION OF MOST SUITABLE METHOD............................................................................................. 37

7 CHAPTER 7.. FAST LAGRANGIAN ANALYSIS OF CONTINUA IN 3 DIMENSIONS (FLAC3D) ...........................................................38

7.1 INTRODUCTION .................................................................................................................................... 38 7.2 THEORETICAL BACKGROUND .............................................................................................................. 38

7.2.1 Formulation of 3D Explicit Finite Difference Model................................................................. 38 7.2.2 Grid Discretization ...................................................................................................................... 39 7.2.3 Numerical Implemetation ............................................................................................................ 39

7.3 MATERIAL MODELS............................................................................................................................. 40 Model.........................................................................................................................................................42 Example Application .................................................................................................................................42

7.4 PROBLEM SOLVING WITH FLAC3D...................................................................................................... 43 7.4.1 The Finite Difference Grid........................................................................................................... 43 7.4.2 Material Properties...................................................................................................................... 43 7.4.3 Boundary and Initial Conditions ................................................................................................. 43

7.5 COMPARISON WITH OTHER MODELS..................................................................................................... 44 7.6 LIMITATIONS OF FLAC3D .................................................................................................................... 44

8 CHAPTER 8 DESCRIPTION OF THE NUMERICAL MODEL .46 8.1 MODELLING GUIDELINES.................................................................................................................... 46 8.2 MODEL GEOMETRY ............................................................................................................................. 49 8.3 BOUNDARY CONDITIONS ..................................................................................................................... 49 8.4 IN-SITU STRESS FIELD.......................................................................................................................... 51

8.5 MECHANICAL PROPERTIES OF ROCK ............................................................................................. 51 8.6 MODELS USED IN ANALYSIS ......................................................................................................... 52

9 CHAPTER 9........... ODELLING RESULTS, COMPARISON AND DISCUSSION.......................................................................................54

9.1 MODELLING RESULTS................................................................................................................... 54 9.1.1 Results of Model 1................................................................................................................. 54 9.1.2 Results of Model 2................................................................................................................. 57 9.1.3 Results of Model 3................................................................................................................. 60 9.1.4 Results of Model 4................................................................................................................. 63

9.2 COMPARISON OF RESULTS OF NUMERICAL MODELS..................................................... 66 9.2.1 Comparison of the Elastic Models (Model 1 & 2) ................................................................ 66 9.2.2 Comparison of the Plastic Models (Models 3 & 4)............................................................... 67 9.2.3 Results of the Elastic & Plastic Models (Model 1 & 3) ........................................................ 67 9.2.4 Results of the Elastic & Plastic Models (Model 2 & 4) ........................................................ 68

9.3 DISCUSSION ABOUT ROCK SUPPORT.............................................................................................. 69 Where, L = bolt length................................................................................................................70 B = span.......................................................................................................................................70

10 CHAPTER 10 CONCLUSIONS & RECOMMENDATIONS.......72

REFERENCES ............................................................................................75

CHAPTER 1

INTRODUCTION

1.1 GENERAL The Hydropower has been developed to an extent in India utilising the clean natural water resources. The country is still facing power defeicit and the peak power demand is high during morning and evening. India has a great hydropower potential estimated as 84044MW at 60% load factor. A large potential of hydropower can be still be exploited from the natural resources available in India. In general, the geology varies considerably throughout India. In the north-west, north and north-east of the country the Himalayas are geologically young active and have complex geological formations. The mountains are full of regional faults and thrust zones. It is quite a challenge to plan, design and execute underground activities in Himalayan region. The geological structure of the Himalayas fall into three broad stratagraphical belts or zones: • The Northern Himalayas or Tibetan zone – composed of marine sedimentary rocks

containing high fossil density. The rocks range from the earliest Palaeozoic to Ecozoic ages.

• The Central Himalayan Zone- consists of the lessor or middle Himalayas composed of mostly of crystalline and metamorphic rocks gneisses and schists with non-fossil bearing sedimentary and Mesozoic deposits

• The outer or sub-Himalayan zone- corresponds to the Siwalik ranges and is mainly composed of tertiary sedimentary river deposits.

Geological and geo-technical investigations of Teesta-V project in Sikkim, India has revealed the presence of phyllite and quartzitic phyllite in the area. Hydropower development in Norway has experienced with various types of similar rocks in Norway. Norwegians have constructed 200 underground power stations and more than 4000 Km of hydropower tunnels in various terrains with innovative new designs and technology resulting in optimum solution for hydropower production. Phyllite and quartizitic phyllite rocks have been encountered in many projects in Norway with comparatively lesser depth of weathered rock. The Norwegian experience in hydropower development could be quite useful in India to evaluate the design in Teesta-V project located in Sikkim, India taking into account the sufficient clearance for weathered rock layer in the underground works.

1

1.2 PROJECT BACKGROUND The proposed Teesta-V project is located in Sikkim State in the North-Eastern part on India. The proposed dam site is located at a distance of 140 Km from the nearest Bagdogra airport and the powerhouse site is located at about 110 Km from Bagdogra airport. The nearest railhead is New Jalpaiguri and Siliguri. (Project Location Map- Drawing No. 01) The proposed project is located in Sikkim, India designed as a run of the river project on river Teesta in the cascade development of the river utilising a gross head of 217m. The project has a catchment area of 4300 square km. The water from the river Teesta is proposed to be diverted through a 97 m high concrete gravity dam and conveyed through a 17.8 km long 9.5m diameter headrace tunnel to an underground powerhouse having an installed capacity of 3*170MW. The project area consists of low grade metasedimentary rocks belonging to Daling group of lesser Himalayas. The dam site vicinity has phyllitic quartzite with thinly bedded Quartzite exposed at certain locations. The headrace tunnel, power house and surge shaft area is represented by low grade sedimentaries with interbedded sequence of Phyllites, Phyllitic Quartzites, fine grained Quartzites and Mylontic Quarzites. The feasibility report of Teesta-V project shows a fully concrete lined headrace tunnel 17.8 Km long and 9.5 m diameter from intake to an open surge shaft and 3 Numbers of 4.7m vertical pressure shafts designed from the surge shaft to the Underground power house comprising of 3 units of 170 MW each. The Power house is built in two main caverns one for machine hall 100 m X 22 m X 46 m and transformer hall of size 82.5 m X 15 m X 25 m.

1.3 OBJECTIVES This study is aimed to design the different alternative tunnel layouts with minimum rock support arrangements with optimum length and optimum number of adits absolutely required for economic construction of headrace tunnel. The alternate tunnelling methodology with mainly using Tunnel boring machines in addition to traditional Drill and blast method with mostly unlined tunnel has been studied based on Norwegian experience. The possibilty of an unlined pressure shaft with a suitable layout has been studied. A compact cavern for locating underground power house with a smaller span has been attempted based on Norwegian experience. The method is based on the literature review and empirical methods of tunnel and cavern design. The results obtained from empirical methods are compared with the Numerical model simulating the ground conditions on the numerical model and check the results.

CHAPTER 2

GEOLOGY AND GEOMORPHOLOGY

2.1 REGIONAL GEOLOGY

The Teesta HE Project, Stage-V is located in South & East Sikkim in India. In the project area low grade meta sedimentary rocks belonging to Daling group of Lesser Himalaya are exposed.In the extreme South the foothills are occupied by Tertiary sediments viz Siwalik group of rocks which comprise alternating sequence of sandstone, siltstone and pebble beds. This group is least deformed and occupies a width of about 2-10kms, trending East-West. The Lesser Himalaya in this region is represented by Younger sedimentaries of Gondwana group in the South. These are overlain by a low-grade metamorphic sequence of Daling-Buxa Group. This group is thrust over the Gondwanas. The Daling-Buxa Groups comprise slaty quartzites & chloritic phyllites etc. (Feasibility Report -Teesta H.E. Project Volume –I) The dam site is located 2 Km downstream of confluence of Teesta river with Dikchu nala. The bedrock is Phyllitic quartzite.

2.2 HEAD RACE TUNNEL The feasibilty Report of the project shows a 17.68 Km.long, 9.5m diameter circular shaped headrace tunnel is envisaged to convey water from Intake to surge shaft.. In the entire alignment the tunnel has three kinks and one curvature near surge shaft. The kinks have been provided keeping in view the requirement of construction adits needed for excavating this long tunnel. The kinks also ensure that, adequate cover is available through out the tunnel length. The main topographical features along this alignment are N-S trending, parallel hill ranges with E-W cross drainages forming deep valleys. Ranchang and Samdong Kholas are the most important drainage channels in this area. The ground profile along the tunnel alignment is low to moderately high, ranging from EL 575M to 1330M. Essentially the low cover reaches are confined to prominent nala crossings where the overall cover is reduced considerably but within safe limit. These stretches may act as avenues of copious water seepage. However, in Rangchang khola, corresponding to chainage about 7550m. The tunnel area has been studied with a backup of systematic geological mapping in 1:25000 scale along the old alignment by GSI and a traverse geological mapping by NHPC along the revised tunnel route. The adit locations have been fixed tentatively and shall be firmed up after the site investigations are accomplished. GSI has also carried out point load

strength index determination on bulk samples collected at 9 locations along the old tunnel alignment marked on ground. The results are tabulated in Table 2.1 below.

Table 2.1 Rock Mass Description Along Tunnel Alignment Samples Nos.

Sample Rock

Type

Point Load Strength (Mpa)

Computed compressive

strength (Mpa)

Orientation

CS1 Quartzose phyllite

4.4 70 Perpendicular to foliation

plane. CS2 Phyllitic

quartzite 4.4 70 Parallel to

foliation plane. CS3 Quartzite 14.1 225 Massive rock. CS4 Quartzose

phyllite 13.4 213 Perpendicular

to foliation. CS5 Quartzose

phyllite 14.1 225 30º with

foliation plane CS6 Phyllite 7.0 113 Perpendicular

to foliation plane.

CS7

Phyllitic quartzite

6.0 95 Perpendicular to foliation

plane. CS8 Soft phyllite

3.9 60 Perpendicular

to foliation plane.

CS9 Fine Grained Quartzite

0.7 10 Perpendicular and parallel

direction with the bedding

plane

The area exposes predominantly low-grade metamorphite comprising phyllitic rock with minor quartzites. Based on the field observation and microscopic study, rocks of this area can be classified into the following units: (i) Phyllites (ii) Phyllitic quartzite (iii) Fine grained quartzite and (iv) Mylonitic quartzite interbedded with thin phyllite bands. The third generation folds (F3) have affected all the previously mentioned planer structures during the later deformation and have mainly produced culmination and depressions. These are open and round hinge upright folds. The general trend of the formation is N-S to NW-SE with dips of 45-50o in E to NE direction. The rocks are

highly jointed with diversity in their orientations. The important joint sets, in order of prominence are mentioned as Table 2.2

Table 2.2 Orientation of Joints Along Tunnel Alignment S.No. STRIKE DIP

1 N-S to N50°W-S50°E 20 to 40° dip towards E to N 40° E

2 E-W 35° to 40° dip towards south/60° to 80°

towards north 3 N 15° E –S 15° W 35° to 65° towards N 75° 4 N45° to 65°E- S45° to 65°W 35° to 85° towards N 25-45°W 5 N70° to 80°E - S70° to 80°W Subvertical to vertical. 6 N50° to 80°W- S50° to 80°E 40° to 60° towards N 10 to 40°E 7 N55° to 65° W- S55° to 65°E 75° towards N 25-35° E to

vertical Shear zones with varying width from 0.5 m to 1.0 m and diverse orientations are observed in the area. The description of these are given below:

Table 2.3 Orientation of Shear Zones S.No. Strike/Trend Dip Location

1 E-W

35° S 1.2km. NE of Mangkha

2

N5°E - S5°W 90° 0 .4km N of Tumin

3 N80°E-S80°W 80° N 10° W 1km NNE of Samdong bazar

4.

N80° W - S80° E 75° N 10° E N of Kambul

5. E-W

15° N 0.5 km SW of intake.

RD wise anticipated rock quality is furnished below. R.D. 0 to 2062m. Highly sheared phyllitic quartzite and phyllite rock is expected at tunnel grade. Two prominent shear zones are expected at ch. 1462m & 1875m respectively. The rock mass fall in poor to fair class.

R.D. 2062 to 3250m Bands of phyllitic quartzite and phyllites are expected. Assessment of RMR & Q values for similar rock in Power house drift has shown values ranging from 48-50 and 24-34 respectively. By & large favourable tunnelling conditions are anticipated in this region.

R.D. 3250 to 8012m Highly folded and sheared bands of phyllitic quartzite and phylllites are anticipated with prominent shear zones at ch. 3250m and 6325m. The rock mass fall in poor to very poor class. Overall tunnelling conditions will be unfavourable and problematic. R.D. 8012 to 16750m. Tunnel would be excavated in predominent phyllites and subordinate phyllitic quartzite. Two prominent shear zones will be encountered at ch. 8950m & ch.11187m respectively. Assessment of RMR & Q values for similar rock in Power house drift has shown values ranging from 58-60 & 24-38 respectively. By and large favourable tunnelling conditions are anticipated except shear zones. The rock mass is expected to fall in fair to good class tunnellling media R.D.16750 to 17687.5m Tunnel would be excavated in phyllites and phyllitic quartzite. The rock mass is expected to fall in poor to fair class. 2.3 SURGE SHAFT AND PRESSURE SHAFTS: A 95m high and 25m dia surge shaft is proposed in hill mass near village Dipudara on Song-Khomdong road. The hill, in which these two structures are located, runs parallel to the river course (N-S) and starts with riverine terrace on the left bank of Teesta river. It rises upto 170m between EL 368M & 538M in the form of rock escarpment, consisting of jointed phyllitic- quartzite, quartzite, and quartzitic-phyllite. The formation trends NNE-SSW with dips of 60o-70o WNW into the hill. Further up, at least up to next 200m,. it is covered by talus and slope wash debris, having a slope angle of 30o to 40o. The overall situation is considered favourable for the stability of hill slope as well as for underground structures. This last hole has shown overburden of just 11.8m at its location, indicating that surge shaft and pressure shafts would be housed within bedrock

A 57m deep drill hole having collar EL 591.87m, has been done for fixing up surge shaft location. The hole has passed through overburden material upto 11.8m. It is followed by fractured and sheared bedrock, comprised of quartzitic phyllite with phyllite interbands. The hole could not touch the tunnel grade due to jamming of the core barrel at 57m depth. The last 3m core also could not be recovered due to the above reasons. The water percolation test suggested low to moderate permeability of overburden media. The water loss between 6- 7.5m depth indicate open voids. The core recovery varies from 33 to 55%, RQD ranges between 3 to 7% due to fractured nature of bed rock. 2.4 POWER HOUSE An underground Power house having three units of 170 MW is proposed on the left bank near Sirwani village. The main Power house cavern is 22m wide, 46m high and 100m long. This cavity is housed within a fairly competent rock unit consisting of predominantly quart-zite with phyllitic bands. The orientation of the longer axis of cavern has been kept askew (25o-30o)to the major discontinuity of foliation plane.

In order to ascertain the rock conditions at power house site, a 110m long exploratory drift was excavated. This drift with invert level at EL 360m is driven in N60oE direction During the field investigations, carried out by NHPC, the drift was extended further up to 140m, then cross cuts were provided for the longer axis of the cavern. The right cross cut was extended for 20m out of 80m proposed length and the left crosscut upto 20m. The details of exploratory drift and its cross cuts are given below. The main drift and its cross cuts, along the excavated length, show wet to dripping water condition. No major overbreaks have been noticed and no support provided except between ch. 105-110m, where cavity formed. The foliation in general is N30oW-S30oE to NW-SE with dips of 30-45o in NE direction. Quartz vein/shear zone/clay seams are mainly oriented along the foliation plane. The right cross cut appears to be more competent with quartzite and quartzitic phyllite constituting the major rock type. The rock mass classification and flat jack tests were performed between 0 to 100m RD of the drift. In addition, 2 nos. of seismic profiles, each 110m long were also laid in the vicinity. Based on investigations done so far, following subsurface picture has emerged: 1) The Power house will be located around 200m inside the hill face, from the left bank of

river, having vertical cover of 170m above its crown. The invert level at about 340 m has been kept 34.5m below the MWL of the river i.e. 374.5m, at this location.

2) The rock unit is well jointed quartzite with subordinate phyllite band. The foliation of

the bed rock orientation in power house area is given below in Table 2.4. 3) Notable seepage may be encountered during excavation. 4) The `Q' system of rock mass classification (Barton) has revealed poor to good rock

having `Q' values varying from 3 to 38.03.The `RMR' system (Bieniawaski) has shown fair to good rock having `RMR' values varying from 45 to 66.

5) The flat jack tests have indicated residual stresses in horizontal direction as 2.18MPa and in vertical direction as 2.26MPa, simulating with hydrostatic stress field condition around the power house cavity. Owing to the shifting of surge shaft location, the power house complex has also been shifted by around 100m towards North.

Table 2.4 Orientation of Joints in Underground Powerhouse Location S.No. STRIKE DIP SPACING NATURE

1 N 30°W-S 30° E to NW-SE

35° to 45° NE - -

2 N40° W-S40°E; 45° N50°E 5 to 30cm Continuous, Rough to undulatory

3 N55°E-S55°W

60° S35° E 15 to 20 cm Rough surface and continuous

4 N85°E-S85°W 60° - S55°E 10 to 30cm Rough to undulatory and

continuous 5 N70° W-S 70° E 55° -S20°W 10 to 30cm Rough and

continuous 6 N20°E-S20°W.

50°-N 70°W 15 to 40cm Undulatory to

rough, not continuous

CHAPTER 3

DESIGN OF THE TUNNEL SYSTEM

3.1 GENERAL The Teesta-V project, in Sikkim, India shows a favourable geological condition for the stability of an underground powerhouse as compared to the surface powerhouse. The optimum solution of a tunnel or an underground powerhouse is largely dependent on the most optimum design and most optimum cost of the tunnel system or the powerhouse cavern. The engineering geological investigations are the most important basis for assuming the design considerations for the tunnel system. The geological conditions in the area are considered for the design of the tunnel system to find the best location, orientation, shape and dimension of the tunnel. Norway is a leading country in using the rock strength for hydropower development after having completed about 3500 Km length of hydropower tunnels and half of the total number of underground power-houses in the world. The experience acquired by the Engineers with these works and using the modern technology has resulted into the most cost-effective solution with the best layout, optimum structural design and effective rock-supports have been developed. These designs have been studied in the following chapters in addition to the standard design procedures. The Feasibility Report of the Teesta-V project has been prepared. The location and orientation has been selected after preliminary geological investigations. Accordingly, the design the headrace tunnel and tailrace tunnel is concrete lined alongwith 3 numbers of vertical concrete lined pressure shafts. Various Empirical methods based on the Norwegian experience for prognosis of work and estimated cost have been applied to find an optimum solution for design of tunnels and pressure shafts. (Nielsen, Theidemann, 1993)

3.2 NORWEGIAN DESIGN APPROACH 3.2.1 Location of the Project The project is located in Sikkim state in North Eastern part of India ( Drawing No. 01). The project lies in the Lesser Himalayas containing mainly phyllite and Quartzitic phyllite. The area indicates shallow to moderate overburden of 5-20m and the rock quality varies from poor to fair rock. A few shear zones are anticipated in the area with width 0.5-1m. The point load strength measurements indicate compressive strengths varying from 60-225 MPa. At certain location it is found to be 10 MPa. The project location is considered good for location of an underground powerhouse and tunnels.

3.2.2 Orientation of Tunnel Axis The length axis of the tunnel and cavern openings at shallow to intermediate depths should be oriented along the bisection line of the maximum intersection angle between the predominant joint directions. Close parallelism to the third or fourth joint directions should be avoided. In evaluating the joint sets, the character of joints with frictional properties, number of joints and dip angles should be considered. The tunnel stability is generally reduced and the overbreak increases gradually when the angle between the tunnel axis and the predominant joint sets becomes smaller than 25-30°. (Nielsen, Theidemann, 1993) The Rose diagram has been developed which shows the Orientation of major joint sets along the tunnel alignment and the orientation of joint sets has been shown in Table 2.2. In the proposed tunnel alignment the joint sets have formation in N-S to NW-SE with dips of 45-50° in E to NE direction. The rocks are highly jointed with diversity in their orientations. The tunnel alignment has been proposed at angle of about N30°W based on the Rose diagram and it has to further checked at the site to suit the topography and to reduce the tunnel stability problems. Fig. 3.1 Rose Diagram for Tunnel Alignment of Teesta- V H.E.

Project

3.2.3 Shape of the Tunnels Circular section gives the hydraulically optimum cross section of a waterway. Also the compressive stresses are distributed evenly along the periphery reducing stress concentrations. Tunnel sections with small curvature radius gives high stress concentrations and anisotropic stress situations. In the underground rock excavations, due to practical restrictions circular sections are only possible with Tunnel Boring Method (TBM). The short tunnel lengths are generally not economical. The best tunnel section hydraulically and practically is the conventional tunnel section with arch roof and vertical side wall. The hydraulically best tunnel section has tunnel height equal to tunnel width. A circular tunnel section for TBM tunnels and conventional drill and blast method has been used for tunnel optimisation. In the Teesta-V project the stress level is moderate but it may be high at certain locations with high stress due to rock cover more than 500m. This may lead to squeezing problems in weak phyllite rock at high stress conditions. It has to be checked separately with numerical models. 3.2.4 Dimensions All categories of rock mass have their own self supporting limits and stability problems in underground excavations. The stability problems of underground caverns and tunnels increase with the increase in excavation span. In Norway most of the caverns have span of excavation varying from 12 to 20m. (Nielsen, Theidemann, 1993)

3.3 ROCK MASS CLASSIFICATION ALONG THE TUNNEL SYSTEM During the Feasibility and preliminary design stage of the project, very little information of rock mass and stress is available. The rock mass classification system can be quite useful at this stage. The classification is used to obtain a general rating of rock mass quality and classification of special rock engineering properties such as drillability and blastability and to estimate the rock supports. Two commonly used methods are: 1. Q- method 2. RMR method 3.7.4 Q- Method The Q- method is based on the determination of tunnelling quality index 'Q- value'. It was developed by NGI ( Norwegian Geotechnical Institute). This method was developed on the basis of evaluation of large number of case histories of underground excavations. The

Q-value determination is based the following six parameters (Barton, N. et al, 1974) and the methodology for determining is enclosed in Appendix 1.03.

SRFJ

xJJx

JRQDQ w

a

r

n

=

Where, RQD = Rock Quality Designation, Jn = Joint Set number Jr = Joint Roughness number Ja = Joint alteration number Jw = Joint water reduction factor SRF = Stress Reduction Factor 3.3.2 RMR Method The Rock Mass Rating (RMR) system has been developed to classify the rock based on various parametrs as shown in Appendix 1.04. The RMR system of Rockmass classification was developed by Bieniawski in 1973 and it involves the six parameters. (Bieniawski 1974) The following six parameters are used to classify a rock mass using the RMR method: 1. Uniaxial strength of rockmass 2. Rock quality designation (RQD) 3. Spacing of discontinuities 4. Condition of discontinuities 5. Ground water condition 6. Orientation of discontinuities

3.6 ROCK MASS RATING ALONG THE ORIGINAL ALIGNMENT Data from the geological investigations is used to assess the rockmass rating along the tunnel system using Q-method and RMR system and the results are presented in the Table 3.1.

Table 3.1 Rock Mass Rating along Tunnel Alignment S. No. RD (m) RMR Q - Value

1. 0 - 2062 30-45 1-10 2 2062 - 3250 48-50 24-34 3 3250- 8012 15-25 0.1-4 4 8012- 16750 58-60 24-38 5 16750 - 17688 30-45 1-10

3.6 DESIGN PRINCIPLES FOR UNLINED HEADRACE TUNNEL / PRESSURE SHAFTS DESIGN:

Stability and water tightness are the main requirements for the unlined headrace tunnel and pressure shaft design. To achieve this objective the location of tunnel system must be selected to satisfy certain geological and topographical restrictions. 3.5.1 Geological Restrictions: The following geological restictions have been looked into to design the unlined tunnel and pressure shaft: 1. High porosity rocks like volcanic and Sanstone : Teesta-V project area comprises

of mainly good quality Phyllite and Quartizitic phyllite. 2. Karstic areas: Geological and geotechnical investigations have revealed that there is

no presence of any calcareous rock types. 3. Heavily jointed rock mass and open interconnecting communicating joints:

Phyllite rock present in the area is of fair quality rock with generally rough and tight joints of average spacing varying from 0.5 to 1m. The permeability tests have revealed low permeability but may have high permeability in certain areas with shear zones and fractured areas.

4. Weakness Zones and faults with unfavourable Orientation: Along the tunnel alignment several main joints and shear zones have been identified which are shown on the rosette diagram enclosed as Fig. 3.1. The orientation of tunnel has been fixed so that the orientation of tunnel is favorable.

5. Impermeable rock layers or clay zones between tunnel/ shaft and the surface: It may cause high pressures to build up in critical locations: Geological and geotechnical investigations have not revealed that any such clay zones are present in this area.

As no major geological problems are anticipated based on above, hence the Teesta-V project area has been studied for unlined pressure shafts and tunnels based on above restictions. 3.5.2 Topographical Restrictions: For the pressure shafts and tunnels the sorrounding rock mass should not be deformed due to water pressure from the tunnel or the shaft. Therefore, the normal stress across all discontinuities in the rock mass must be higher than the water pressure in the unlined tunnel/shaft. Otherwise the hydraulic jacking of the discontinuities may take place. Instantaneous closure pressure in hydraulic jacking (Pisi) is the most direct parameter in connection with the evaluation of the possibilities for water leakage.

Table 3.1 Evaluation for Hydraulic Jacking in Tunnels/Pressure shafts Closure pressure in hydraulic jacking (Pisi) 2.13 MPa Maximum static pressure at the junction between pressure shaft/tunnel ( Ptsw)

0.3 MPa

Maximum static pressure at the junction between pressure shaft / penstock

2.1 MPa

Factor of safety against leakage Pisi / Ptsw 1.0 The Teesta-V area falls under requirements of topography restriction with a factor of safety of 1.0. This indicates that considering a facor of safety of 1.5 only the upper portion of the length of the pressure shaft upto a head of about 130m i.e. maximum satic pressure of 1.4 Mpa can be kept as unlined tunnel and rest of the portion may be lined with concrete provided the geological conditions are favourable. It may further be analysed using Numerical modelling of the pressure shaft area.

3.6 OPTIMISATION OF WATERWAYS FOR UNLINED TUNNELS Cross sectional areas of the tunnels and shafts are optimised using the cost of construction and the loss of power. The criteria for the calculation of the optimum cross-sectional area is to find the area such that the sum of the cost of construction and the capacity loss is minimum. Head loss in the tunnels is used to calculate the capacity loss, which is expressed by Manning's formula:

LRMA

q

h ×

= 3/42

2

Then, capacity loss is given by:

LRM

Aq

qC ××

×××= 3/42

2

81.9 η

Where, h- Head loss in Tunnel, q- Discharge in m3/s, A- Cross-sectional area of the tunnel L- Total tunnel length, M- Manning's constant, R- Hydraulic radius of tunnel, η - Efficiency of the plant

Waterways optimization has been carried out for various alternatives such as lined and unlined tunnels and on original alignment in the Feasibility report and the the modified shorter alignment. The tunnel optimisation study is enclosed in the Appendix-1.1 for concrete lined tunnel with the original alignment and Appendix 1.2 for concrete lined tunnel with the alignment No. 2. The Appendix 1.3 shows the optimisation study for unlined tunnel with the original alignment and Appendix 1.4 shows tunnel optimisation studies carried for unlined tunnel with the shorter alignment No. 2 .

3.7 SELECTION OF AN OPTIMUM LAYOUT Depending upon the topography of the area and the rock mass quality of a particular project, different alternative layouts could be studied. Unlined headrace tunnels, unlined pressure shafts with an air cushion surge chamber are the new design concepts which are commonly being used in the Norwegian hydropower industry. These new concepts have reduced the lengths of tunnels and steel requirements for construction saving time and total cost of the project considerably. The various option have been studied to get the optimum solution for tunnel costs and the construction time could be reduced so as to get the economic return is maximum from the project. The study for optimisation of lined and unlined tunnels for various tunnel cross-sectional areas with cost of construction and energy loss is enclosed in Appendix 1.1 to 1.4. Due to limited data available for cost of construction materials and value of firm power in the area, the values from similar Norwegian projects have been considered for analysis. The optimisation studies indicate that 80m2 of tunnel cross-sectional area for lined tunnel and 200m2 of unlined tunnel cross-sectional area will be the most optimum section. Considering the rock quality and the topography of the Teesta-V project area, the following alternatives have been studied: Option 1: The headrace tunnel with five numbers of adits and tail race tunnel alongwith a vertical pressure shaft and an open surge chamber.as shown in drawing No. 06 and Appendix 1.01. This option has been studied with drill and blast method from all 5 adits as indicated in the Feasibilty Report (This arrangement has been selected at the feasibility design level with concrete lined tunnels.) Total lengths of adits = 1.54 Km Head race tunnel length = 17.8 Km Tail race tunnel length = 0.145 Km The tunnel drill and blast prognosis has been done in Appendix 2.1 based on the limited data available regarding the rock in the area. A tunnel of 80m2 of cross-sectional area in the phyllite rock considered as poor blastabity and poor drillability may give an advance rate of about 40m/week with drilling jumbos of 3.3 drilled round length using empirical

relations based on Norwegian experience. (Ref: Drill and Blast Prognosis, NTNU) the drill and blast costs for excavation of 80m2 has been done in Appendix 2.2 is about NOK 12000/m i.e.US$1300/m. 3.7.2 Option 2 The head race tunnels 2 numbers unlined each 9m diameter have been studied with 3 Nos. of adits on revised alignment No.2 as shown in Drawing No. 06 to reduce the lengths of adits and reduce the construction time and cost. The alternative construction methodology using Drill and blast method in parts of the tunnel with poor rock conditions and good rock area with Tunnel Boring Machine (TBM) has been studied to reduce the cost and time of construction. The good rock in 8.5 Km length is proposed to be bored with TBM and the remaining 6.94 Km in fair to poor rock is proposed with Drill and Blast method. Total lengths of adits = 0.940 Km Head race tunnel length = 15.44 Km Tail race tunnel length = 0.145 Km The TBM tunnel alternative has been considered for part of tunnel as enclosed in Appendix 2.3. Using two numbers of tunnels each 9m diameter for 8.5 Km length of tunnel, the weekly advance rate for 100hours/week is 92m/week.The cost of excavation using TBM is about NOK 9000/m i.e. about US$ 1000/m. The same tunnelling done as full length of 15.44Km with TBM as enclosed in Appendix 2.5 may cost slightly higher i.e. US$1050/m. The proposed construction schedule for the option is shown in Appendix 3.1. 3.7.3 Option 3: The head race tunnels 3 numbers, unlined each 7 m diameter one for each unit of 170 MW have been planned with 3 Nos. of adits to reduce the lengths of adits and reduce the time of construction and cost. The construction methodology using Drill and blast method in the part of tunnel with poor rock conditions and Tunnel Boring Machine (TBM) in remaining part of the proposed tunnel has been studied to reduce the cost and time of construction. The good rock in 8.5 Km length of tunnel is proposed to be bored with TBM and the remaining 6.94 Km in fair to poor rock is proposed with drill and Blast method. Total lengths of adits = 0.940 Km Head race tunnel length = 15.44 Km Tail race tunnel length = 0.145 Km The prognosis of 7 m TBM and drilll and blast have been carried and are enclosed as Appendix 2.6 and 2.7 Appendix 2.4 with 3 Numbers 7m diameter tunnels for 8.5 Km length alongwith drill and blast for remaining lengths indicates that the cost of tunnelling is about NOK 9000/m ie US$1000/m. The advantage with this method can be evaluated

based on revenue due to commissioning of 1 unit 2 years prior to the other alternatives which can generate return on the investment for additional units. 3.7.4 Option 4: The head race tunnels 2 numbers each 9m diameter have been planned with 2 Nos. of adits to reduce the lengths of adits. The complete tunnel has been proposed to be excavated using Tunnel Boring Machine (TBM) and the main features of this option are: Total lengths of adits = 0.375 Km Head race tunnel length = 15.44 Km Tail race tunnel length = 0.145 Km This option reduces the length of adits considerably but it may have the problem in boring with TBM in weak rock areas. This option has the advantage of completing one tunnel and commissioning one unit quite early so that the return on investment from this unit can be further utilised for investment in construction of additional units. It requires planning for treatment of weak zones in advance and selection of a proper TBM after having done the detalied geological investigation of the tunnel area and samples of rock have been tested in the laboratory to predict the performance of TBM. The TBM should be used based on the technical and economic feasibilty of the option. 3.7.5 Selection of Best Option The various above options have been studied for preparation of the construction schedule of the project in Appendix 3.1. This option has been considered as the best option based on the construction schedule and other field data available from the area. It indiactes that the option 3 which is proposed to have 2 unlined tunnels of 9.0 m diameter and toal length of 15.44 Km with 8.50 Km length excavated with TBM and remaining 6.94 Km with drill and blast method has been considered as the best option. The nature of rock in the tunnel area is mostly fair to good with only small areas of poor rocks. It can be considered as suitable for Tunnel Boring machine in fair to good rocks and drill and blast can be used in Poor rock location. The Option 3 with 2 numbers of tunnels each 9m diameter shows that 1 units of 170 MW could be started after 5 years of construction and the second and third units after 2 years after completing the first tunnel. This can further improve the econmic viabilty of the project.

CHAPTER 4

DESIGN OF AN UNDERGROUND POWER HOUSE

4.1 GENERAL Locating an underground powerhouse in a certain area gives an additional degree of freedom as compared to the surface powerhouse. An Underground power station has a characteristic feature to reduce the length of waterway such as tunnels and pressure shaft and the strength of rock is used to support the tunnel system without concrete or steel lining. An Underground power stations is considerably much economical than a comparable surface power plant. The underground powerhouse is planned to have a compact and economic size at a suitable location, with good orientation, optimum size and the layout of the cavern suitable with respect to the geological conditions of the area.

4.2 NORWEGIAN DESIGN APPROACH The general procedure followed in the Norwegian design of underground power stations is divided into 4 stages. (Nilsen, Thidemann, 1993) • A location representing the optimum rock stability is selected for an underground

powerhouse. • The Length axis of a powerhouse opening is oriented in such a direction to give the

minimum stability problems and the minimum over break. • The opening is shaped in accordance with the mechanical properties and joints of the rock

mass as well as the local stress condition, prevailing in the rockmass. • The various parts of the underground structure are dimensioned to give the most economic

results. 4.2.1 Location Investigations during the feasibility study are done to find the best location of the underground powerhouse. The location of the powerhouse is the most important as it depends upon the quality of rock mass and the other geological and topographical conditions, the decision regarding the location of the powerhouse is the most important. Results of the investigations of the area are studied to avoid unfavourable rocks and to estimate the depth of weathering to define a minimum rock cover for shallow or moderately deep-seated caverns. The stress measurements and other detailed geological investigations are required in addition for deep-seated caverns. It is necessary to pay attention to the orientation and location of any major weakness zones and the best location is the one, which does not intersect with any of the weakness zones. If, the weakness zone can not be avoided, then the crossing should be made as short as possible. Steep dipping discontinuities will influence the roof stability of the cavern.

Teesta-V project is located in an area consisting of fair to good quality Phyllite and quartizitic phyllites. Average compressive strengths values vary from 60-220 Mpa. The depth of weathered rock is about 5-20m. The rock cover over the power house is about 165 m. At this depth the high rock stresses cannot be expected. But this area is in the tectonic zone and in situ stresses measurements are important to evaluate the actual stress conditions around the proposed cavern. The flat jack tests have been done in the power house area and it reveals that the residual stress in the horizontal direction is 2.18 Mpa and in the vertical direction is 2.26 Mpa.

The maximum tangential stress is calculated as σt(max) = (3σ1-σ3 ) = 4.6 Mpa The minimum tangential stress is calculated as σt(min) = (3σ3-σ1) = 4.28 Mpa Therefore, it is expected that there will not be any stress induced stability problem around the cavern due to the prevailing moderate in-situ stress. The location of the cavern may be selected in such a manner to avoid any weakness zone intersections in the area. 4.2.2 Orientation of Length axis: The most important criteria for orientation of a cavern is normally to orient the most prominent joint set based on orientation and magnitude of the horizontal stresses. Based on the directions of joint and rock stresses, it is possible to give a proper orientation of the cavern to reduce the stability problems and to minimise over break. This requires a comprehensive joint mapping and plotting the observations in a joint diagram like joint rosette. For the openings located at shallow and intermediate depths (low stresses) the length axis should be oriented along the bisection of the maximum intersection angle between the dominant joint directions while close parallelism to the third and fourth joint sets directions should be avoided. In this case, the number of joints are not mentioned but based on the reducing prominence of joints and strike and dip of joints, the orientation of cavern is decided. It is important to have an angle of minimum 25° to steeply dipping smooth joints for the long and high walls of a cavern. In case of high rock stresses, it would be necessary to take into account the direction of principle stresses in addition to the above requirements. Direction of the major principal stress perpendicular to the longitudinal axis, may increase the rock bursting along the roof of the cavern. High anisotropic stresses results in high tangential stresses around an opening. Therefore, the area of tunnel contour tangential to the plane through which the major and intermediate principal stresses are exposed may face the rock stability problems. The most stable orientation is obtained when the length axis of the underground opening makes an angle of 15-30° to the horizontal projection of major principal stress. If the direction of the principal stress is close to the direction of bedding or foliation in highly anisotropic rocks, the length axis of the opening should be oriented with an angle as large as possible relative to the strike of foliation plane. In such situations, 35° should be regarded as absolute minimum.

The joint rosette of rocks in the Teesta-V power house cavern location shows that the length axis of the proposed cavern should be located in North direction which is considered as the best location of the cavern in the area.

Fig 4.1 Rose Diagram for Joints at Underground Powerhouse Location of Teesta –V H.E. Project

4.2.3 Shape The cross- section of the cavern is given a shape to adjust the mechanical properties of the rock mass, joints and rock stresses. The stability of rock depends on the shear strength of the discontinuities and which in turn depends upon the rock stresses. Therefore, based on the rock conditions of the area, openings should be made as different shapes like flat or deep arch roof, vertical or curved walls or asymmetric shape. Intruding corners and edges should be avoided to eliminate destressed areas. If the stress level is too high or anisotropic, it is advantageous to avoid shapes with small curvatures to reduce stress concentration. But in high stresses or too anisotropic stress situations small curvatures are purposely used with asymmetric shapes to concentrate the stresses only in portion of the contour thus reducing the area to be supported. The design concept based on the two different principles is shown schematically on Fig. 4.1. Fig. 4.1 Design Principles for Underground Openings in Rocks at Varying Stress levels and

with varying directions the Major Principal Stresses and Normal to Length axis (Broch,Olsen , 1982)

The figure implies that the stable situation for an opening with simple shape and high walls is obtained when the opening in a rock mass is dominated by moderate horizontal stresses and the length axis of the opening is oriented normally to the direction of the principle stress. The proposed cavern area is located in the rock mass with moderate stresses and with the major principal stress almost perpendicular to the length axis of the cavern. Hence it is possible to give high walls with arch shape roof.

4.2.4 Dimensions Dimensioning of an underground opening is based on the detailed calculations is not very common in Norway. It is due to the main problem of obtaining the reliable parameters for the material and variation of rock mass properties in all the three directions. (Broch, E., et al, 1995) Stability problems in underground opening normally increase with the increase in span. The increase in span of cavern increase the height of the arch to avoid small curvature shapes and then more discontinuities start to cross the openings increasing the stability problems. The orientation of joints may also greatly effect the span of the opening. Horizontal bedding or joints influence unfavourably on the span of the arch. When it needs to increase the volume of the opening, extension should be achieved by expanding the opening along the length axis. Span is the critical dimension for a cavern. For a given rock quality and powerhouse location, the span of the cavern mainly governs the rock support measures required for the cavern. This is illustrated in Fig. 4.2

Fig. 4.2 Rock Support Measures as a function of Cavern width and Rock mass Quality (Barton et. al, 1974)

According to the Norwegian practise, arch height is taken approximately 1/5th of the span. The span of the proposed powerhouse is 22m and hence the arch height should be about 4.4m.

4.3 OPTIMISING OF CAVERN INTERNAL SPACE The internal space requirement is determined by the size of machines to be fixed, ventilation and communication requirements and space for repair and service etc. Development in the turbine manufacturing has achieved compact generators and turbines of high capacities reducing the inner space requirements and thus reducing the volume of rock excavation per MW installed capacity. Therefore, it is advantageous to used bigger capacity generators and turbines instead of smaller units. But, this criteria is overruled by the hydrological and operational requirements. Vertical axis turbines require high narrow caverns, which are preferable to wide shallow caverns for horizontal turbines.

Transformer arrangement within the cavern can greatly affect the space requirement in a power station. Different transformer arrangements are considered for optimising cost of high voltage cables and safety of the power-house. Power stations with Francis turbines could be designed with one shaft for each generating unit to remove both inlet valve and turbine runner from the same shaft. Erection space for the machine hall, facilities for working staff and auxiliary space for repair and maintenance and ventilation are other factors governing the space requirements in a power station. In Norway, for a project of moderate capacity cavern volume of 250 cubic metres per MW installed capacity is the normal volume excluding rock excavation for access tunnels, cable shafts, tailrace etc. but including transformer cavern volume. The proposed Teesta-V project has been designed with 3 units of 170 MW each with vertical axis Francis turbines. The selection of unit size is based on the hydrological condition and power demand in the area. The same capacity of all units reduces the cost of spares and space required for storage of spare parts. One shaft can be used to remove inlet valve and runner.

4.4 SPAN OF THE CAVERN The generator enclosure including the wall thickness, hatch for main inlet valve, passage of 2m on each side of the generator governs the minimum span of the cavern. But, the different arrangements of the transformer layouts will finally decide the span of the power cavern. Span of the proposed cavern is 22m. Calculations indicate that the similar powerhouse can be placed in a cavern with a span of 18 to 20 m. Therefore, all the possibilities to reduce the span of the cavern are studied. Transformer Layouts: The transformer arrangements and its location greatly affect total layout of the power house. Mainly the transformers are placed in two different ways: 1. Transformers are placed in a separate cavern outside the main cavern. 2. Transformers are placed in the same cavern. In the first case the transformers are placed in a separate cavern. In most cases this cavern is combined with the tail-race gate. This layout increases the distance between the transformer and the generator and hence increases the length of the high voltage cables. The distance is governed by the prevailing geological conditions in the area and according to the Norwegian experience, a minimum distance between the two should be about the height of the small cavern. This option will reduce the span of the cavern and it tends to reduce the rock stability problems in the main cavern, but it increases the stresses on the pillar between the two caverns. This layout is preferable with respect to the possible transformer explosions.

In one cavern solution transformers are placed in the main cavern, but three different layouts are commonly used: • Transformers placed on the machine hall floor by the side of the generators. • Transformers placed in an extension of a machine hall. • Transformers placed between the generating units. In the power house cavern, when transformers are placed by the side of the generators on the machine hall floor, it makes the connection length shorter between the generators and transformers and also the connection length does not depend upon the number of units. However, this arrangement requires an increase in span and it is not suitable for the areas with high rock stresses. Possible damage with a transformer explosion is the main drawback in this arrangement A small number of transformers can be placed in an extension of the machine hall. This layout requires long length of high voltage cables between the generators and transformers. The generators placed near the entrance have further added risk of smoke at the entrance in case of the transformer explosion. This arrangement has the advantage of minimum cavern span and is the most suitable in areas with high rock stresses. The third arrangement has the transformers placed in the enclosures between the generators on the turbine floor. Width of the cavern does not affect by this solution but the length of the cavern will be increased. In most cases, this solution gives the most compact arrangement regarding the excavation volume per MW of the installed capacity. The shorter distances between the generators and transformers demands shorter lengths of high voltage cables. Special considerations should be made for the possible consequences of transformer explosion. The cover of the transformer should be fitted with the anchor bolts and pressure relief openings should be provided to avoid any interference with other machines. This arrangement involves more complicated civil works than the layout with a separate transformer cavern. Escape Routes The safety regulations require at least two alternative escape routes in an emergency. Therefore, the power house is provided with two stairways from bottom up to the machine hall. Shafts Shafts should be provided to transport equipment from bottom to the machine hall floor during erection and maintenance. Powerhouses with Francis turbines, main inlet valve and turbine runner are lifted through the same shaft located over the inlet valve. But, in the power house with Pelton turbine units, each unit is provided with two shafts one for inlet valve and the other for turbine runner etc. The second shaft is generally provided at the side of the turbine on the downstream side. Workshops and Repair facilities The modern generators and other equipments are very complex and repairing them at the power station can be very difficult and it may have to be sent back to the manufacturer. But, the repair

of the runners etc. is normally done on the turbine pit. Therefore, the workshop is necessary for other small repairs. Control rooms and Social rooms For remote controlled hydropower stations the staff attends only for periodical inspection once a day and hence large and well furnished control rooms and social rooms are not necessary. In case of a power house which is designed for both manual and remote controlled operations these facilities should be provided as per standard practise. Ventilation Fresh air supply for the ventilation purposes is conducted through cable tunnel or through auxiliary construction adit. Exhaust air is normally expelled through the access tunnel. Certain sensitive electronic equipments demand special environment as they are vulnerable to high humidity and temperatures. Therefore, air conditioning machines are generally fixed at places with such equipments.

4.5 DIMENSIONING OF THE MACHINE HALL 4.5.1 Height The height of the machine hall basically depends on the way main entrance is laid out. Two commonly used arrangements are: 1. The main entrance to the machine hall on a gable wall. 2. The main entrance to the machine hall on a side wall. If the main entrance to the machine hall is through the gable wall, then the height from the floor to the ceiling is governed by the height of the principal hoist including the height of hook and yoke placed at the highest elevation and the height of the highest component to be lifted including its lifting accessories and a tolerance of about 500 mm from the machine hall floor. Normally, the rotor assembly with shafts or main transformer will decide the height, the junction between turbine part and generator part of the shaft should be placed at the highest possible elevation. If the main access to the machine hall is through the side-walls, all transports have to pass under the crane supporting beam. In this case, the minimum height between the floor and ceiling of the machine hall is influenced by the height of the beam, yoke height and the necessary tolerance of about 500mm. The height above the ceiling depends upon the support measures and roof design. The Norwegian practise is to keep the arch height approximately 1/5th span of the cavern, if the stress directions and joints are properly oriented.

4.5.2 Span The span of the machine hall is decided based on the size of the substructure and the design of the crane supporting beams of the machine hall. If the crane support is erected on the columns, span of the cavern will be increased by the width of the columns. In Norway the practise is to support the crane beam by long steel bars in the rock and reduce the cavern width. 4.5.3 Erection Space Usually the shaft and rotor are assembled on the machine hall floor and hence necessary space should be provided on machine hall. Poles are also stored on the machine hall floor temporarily. Correct planning of transport of equipment and parts into the cavern is important to reduce unnecessary storing space in the machine hall.

Fig. 4.3 Different Locations of Transformers (Nakins et al, 1984)

4.6 DIFFERENTPOWERHOUSE LAYOUT DESIGNS 4.6.1 OPTION 1: The transformers are located in the main cavern by the side of the generating units. Span of the cavern = 28m Length of the cavern = 100.5m Height of the cavern = 49 m Total excavation volume of the cavern =137,900m3 Length of the transformer cables =90m 4.6.2 OPTION 2: The transformers are located in the main cavern in between the generating units. Span of the cavern = 18m Length of the cavern = 130m Height of the cavern =49m Total excavation volume of the cavern =114700m3 Length of the transformer cables =75m 4.6.3 OPTION 3: The transformers are located in the main cavern as an extension of the machine hall. Span of the cavern = 18m Length of the cavern = 130m Height of the cavern = 49m Total excavation volume of the cavern =114700m3 Length of the transformer cables =180m 4.6.4 OPTION 4: The transformers are located in a separate cavern close to the main cavern. This layout has been proposed in the proposed design. Span of the cavern = 22m Length of the cavern = 100.5m Height of the cavern = 49m Size of transformer cavern = 82.5 X 15m X 25m Total excavation volume of both caverns =140,000m3 Length of the transformer cables =210m

CHAPTER 5

DESIGN OF ROCK SUPPORTS

5.1 GENERAL APPROACH TO THE DESIGN OF ROCK SUPPORTS Norway has about 400 hydropower stations, out of which about 50% are underground power stations. These hydropower schemes include about 4000 Km of tunnels. Less than 5% of the hydropower tunnels are lined with concrete. The rockbolts and shotcrete have been used for rock support only when strictly required. Local rock falls in the order of a few m3 in the unlined head race tunnel and tail race tunnel during operation has been accepted by the owners. The philosophy for this that the minor blocks of rocks on the tunnel floor do not reduce the water carrying capacity of an unlined tunnel with a flow velocity of about 1m/sec. (Broch,E.,et al, 1996).

5.2 ROCK SUPPORT METHODS The principal objective in the design of an underground excavation support design is to assist the rock mass to support itself. The basic philosophy in tunnelling and rock engineering is that the extent of rock supports should reflect the actual rock condition. During the excavation, at the working face, temporary supports are installed to ensure safe working conditions. Generally rockbolts and shotcrete are used for it. The permanent supports are generally installed after the excavation, behind the working face. The temporary supports will form a part of the total support requirements. The most commonly used rock support methods are: 1. Rock bolting : Spot bolting of individual unstable blocks Systematic bolting of a section of the tunnel/ cavern 2. Shotcreting : Plain shotcrete Fibre reinforced shotcrete 3. Concrete Lining : Plain concrete Steel reinforced concrete 4. Steel Ribs : Steel support as ribs 5. Grouting : Pre-grouting Post-grouting 5.2.1 Rock Bolting The tensioned rockbolts are normally used for immediate support at the working face. These bolts are anchored mechanically or resin anchored. In corrosive environment, grouted bolts are used. These bolts are generally tensioned to 25-50% of their yield strength, except under rockburst conditions. In rockburst situations, a minimum tensioning is applied and large steel plates are used to avoid crushing of the surrounding rock. For. Rockbolting behind the working

face, untensioned grouted bolts are commonly used. Generally, the effect of rockbolting as support for rockmass will depend to a very large extent on the proper direction and the depth of penetration beyond the discontinuity. 5.2.2 Shotcreting Recently, the shotcreting is the most commonly used and well-recognised method for supporting rock mass in tunnels and caverns. Use of the fibre-reinforced shotcrete improves physical properties of shotcete such as shear and flexural strength, durability, reduction of shrinkage cracks etc. considerably. Adding steel fibres about 1% of the shotcrete volume, it increases the load carrying capacity of a 50 MPa shotcrete slab by about 85% and its ductility as much as 20 times the original value. (Nielsen, Theidemann, 1993) Accelerators allow the shotcrete to achieve early high strength, preventing sagging and sloughing of shotcrete during application, thus reducing rebound and increasing the plasticity of the mix. Micro silica in the mix improves the mix properties of the shotcrete and makes the mix workable. Also, it improves its frost resistance. Shotcrete is primarily applied in heavily jointed rock mass as an immediate support with a thickness of about 50mm. In case of heavy rock spalling shotcrete is supplemented with rockbolts and used as a permanent support. Normally this combination has the capacity to replace concrete lining, that was earlier considered as the only option available in such rocks. Shotcrete can be recommended in the areas with weakness zones and faults provided water seepage and swelling clay material is not present in the rock mass. If the leakage is not very high, drainage may be installed along with the shotcrete. 5.2.3 Concrete Lining Concrete lining is now used only in certain areas with exceptionally poor rocks like weakness zones with active smectite. In the presence of swelling smectite, a cast-in-situ concrete lining has the advantage over shotcrete, as it is subjected to a lower swelling pressure due to the incomplete filling of concrete against the crown of the support and shrinkage of concrete. Concrete lining is normally designed with a minimum thickness of 30 cm. The cost of concrete lined tunnels may be as high as 3 to 5 times the cost of unlined tunnels. 5.2.4 Steel Ribs The steel ribs are used at certain locations having very poor rock and squeezing ground conditions. The steel ribs can be placed at a fast rate. Steel ribs cab also be used with shotcrete and rock bolts at certain areas. 5.2.5 Grouting Grouting is very rarely used in the tunnels or caverns except in cases of high water inflow or tunnelling is very difficult or there is a risk of washing out of in-filled materials. Pre-grouting

can be used in major faults and weakness zones to improve the strength of rock and seepage of water in the rock. Pre-grouting (ahead of the face) normally gives better results than post-grouting (behind the face). The grouting is carried out around the concrete plug in pressure tunnel and pressure shaft extensively.

5.3 EVALUATION OF ROCK SUPPORTS The following approaches are used today for the evaluation of rock supports: 1. Empirical Methods : Based on past experience Based on the classification system 2. Analytical Methods : Based on Limit Equilibrium Analysis Based on Numerical Analysis Traditionally, the Norwegian practise is to make the final decision regarding the type and extent of supports during excavation. It is mostly based on the past experience. The data regarding the rock mass properties is collected during the various stages of investigation. It is the main input for evaluating the type, extent and performance of rock supports. Other factors like cost, time consumption, availability and previous experience are also considered for deciding about the rock support. For analytical methods, the main problem is to obtain the reliable input parameters like boundary stress conditions and other mechanical properties of the rock mass. The results from a numerical analysis can never be more reliable than the input parameters. Therefore, the uncertainties and restrictions are connected with the most sophisticated methods of rock support evaluation. The Norwegian trend, however, is to use these sophisticated methods as a supplement to the empirical methods, particularly in poor rock conditions and complex structures such as large span caverns for which not much of the empirical data is available. Rock mass classification systems like Q- method and RMR system have introduced support evaluation mechanics that have been developed after having studied a large number of case histories. The support design is based on the experience, while the final decision is taken during excavation. The classification system gives an initial estimate of support requirements for the Pre-construction stage support evaluation of a cavern in a partcular nature of rock mass. Due to the uncertainties in assigning the rating values for certain parameters, a range for the value is assigned to them initially. Then, the average value of Q or RMR can be used for choosing a basic support system while the range gives an indication of the possible adjustment during excavation. Evaluation systems like rock mass classification systems are subjective and these systems reflect current and past practise that may have been influenced by the local practise and local

geological features and hence, it may not constitute the optimum design methodology. In recognition of above limitations, the following procedures can be used: ◊ ◊ ◊ ◊ ◊

Using the recently developed version of the classification system Use a range of values for uncertain parameters Compare with more than one system Attempt to verify design with modelling and field monitoring Update recommendations as and when information is available.

In this study, based on the limited data available, Q-method is applied and RMR method is used for comparison.

5.4 SUPPORT DESIGN FOR TUNNEL SYSTEM USING Q- METHOD The Q- method has been applied for design supports for headrace tunnel and the results are shown in the table below:

Table 5.1 Rock supports for Head race tunnel Chainage (m)

Description Q- value Supports Recommended

Comments

0-2062 Poor to fair 1-10 systematic bolts and shotcrete 50-90 mm thick

Shear-zones- 2 Nos. –special Treatment reqd.

2062-3250

Fair to Good 24-34 Spot bolting -

3250-8012

Very Poor to Poor

0.1-4 Fibre reinforced shotcrete 100-150mm and spotbolting

Shear-zones- 2 Nos. –special Treatment reqd

8012-16750

Fair to Good 24-38 Spot bolting Shear-zones- 2 Nos. –special Treatment reqd

16750-17690

Poor to fair 1-10 systematic bolts and shotcrete 50-90 mm thick

-

Table 5.2 Summary of the Support Design

Q- value Support Type Percentage of tunnel length 24-38 Spot bolting 56% 1-10 Systematic rockbolts and shotcrete 50-90 mm

thick 17%


22%

0.1-0.4 Concrete lined in shear zones and swelling or squeezing zones

5%

5.5 SUPPORT DESIGN FOR PRESSURE SHAFT USING Q- METHOD The Q- method has been applied for design supports for pressure shaft and the results are shown in the table below:

Table 5.3 Rock supports for Pressure shaft Depth (m) Description Q- value Supports Recommended Comments 0-6m Overburden 0.1-1 Fibre reinforced shotcrete

100-150mm and spotbolting -

6-7.5m Open voids 0.1-0.4 Concrete lined in shear zones and swelling or squeezing zones

Shear-zones- 2 Nos. -special Treatment reqd

7.5-11.8m Overburden 0.1-1 Fibre reinforced shotcrete 100-150mm and spotbolting

-

11.8-21m Phyllite Rock

1-10 systematic bolts and shotcrete 50-90 mm thick

-

21-24m Shear zone 0.1-0.4 Concrete lined in shear zones and swelling or squeezing zones


24-48m Phyllite Rock


-

48-51m Shear zone 0.1-0.4 Concrete lined in shear zones and swelling or squeezing zones


51-120m Phyllite Rock


-


Q- value Support Type Percentage of shaft length 0.1-0.4 Concrete lined 15% 0.1-1 Fibre reinforced shotcrete

100-150mm and spotbolting 10%


75%

5.3 SUPPORT DESIGN FOR POWER HOUSE CAVERN USING Q- METHOD The Q- method has been applied for design supports for powerhouse cavern and the results are shown in the table below:

Table 5.5 Rock supports for Power house cavern

Length (m)

Description Q- value Supports Recommended

Comments

70m Fair to Good 3-38 Spot bolting Grouting required for seepage

20m Poor to Fair 1-10 Systematic bolts and shotcrete 50-90 mm thick

Grouting required for seepage

10m Very Poor to Poor


Grouting required for seepage


Q- value Support Type Percentage of cavern length 3-38 Spot bolting 70% 1-10 systematic bolts and

shotcrete 50-90 mm thick 20%


10%

The above support design done using the Empirical formulae are required to be checked up with the numerical modelling and it has been done in the subsequent chapters using FLAC 3D numerical modelling of power house cavern.

CHAPTER 6

NUMERICAL METHODS IN ROCK ENGINEERING __________________________________________________________________

6.1 GENERAL

Numerical methods of analysis are now widely used in the field of rock engineering. Numerical methods represent the most versatile and complex group of computational methods used in the field of rock engineering. The purpose of carrying out numerical analysis varies. It can be used to carry out qualitative analysis to understand the behaviour of rockmass or the failure mechanisms.Parametric analysis and Sensitivity analysis can be carried out for comparison and better qualitative assesment. Quantitative results are expressed in absolute values.Numerical anlaysis is often done to determine the deformations expected and to study the effect of rock support. The numerical methods used in the field of rock engineering are listed in Table 4.1. ( Sinha, R., S., 1989) Table 6.1 Numerical Methods in rock engineering ( Sinha, R., S., 1989)

NUMERICAL METHODS

CONTINUUM MODEL

DISCONTINUUM MODEL

SUBGRADE REACTION MODEL

FINITE ELEMENTMETHOD

BOUNDARYELEMENTMETHOD

DISCRETEELEMENTMETHOD

FINITEDIFFERENCE

METHOD

BEAMELEMENTMETHOD

A brief review of these numerical methods is given in the following section. 6.2.1 Finite Element Method (FEM) In this method the subsurface is predominantly modeled as a continuum. Discontinuties can be modeled individually. The problem domain is discretized into a limited number of elements that are connected at nodal points. Each element is finite. The stress-strain relationship is defined by an appropriate constitutive law. The stress,strain and deformtion to be analysed are caused by change in the subsurface condition ( for example excavation). Stress , strain and deformation induced in one element impacts the behavior of its neighbouring elements and so forth.The analysis is performed by solving the equation matrix that models the mesh. The finite element method is well suited to solving problems involving hetrogeneous or non linear material properties, since each element explicitly models the response of its contained material. The finite element uses implicit solution technique and requires large computing capacity. With the advancement in computer technology and higher speed of calculation the finite element method has become very useful these days.A representative program for FEM is ABAQUS. 6.2.2 Finite Difference Method (FDM) The method is similar to the finite element method in that the subsurface is modeled as a continuum that is divided into a number of elements which are interconnected at the nodes. The primary difference lies in the approach used to solve the unknown parameters.The Finite difference method is based on the explicit approach.The explicit method builds on the idea that for a small enough time step,a disturbance at a given mesh point is experienced only by its immediate neighbours. This implies that the time step is smaller than the time that the disturbance takes to propogate between two adjacent points.Initially conceived as a dynamic computation approach the FDM can be used to solve static problems by damping the dynamic solution. A representative package for the Finite difference method is FLAC3D . FLAC3D is a three dimensional explicit finite difference program for engineering mechanics computation.In this study FLAC3D package has been used and is described in detail in the next chapter. The advantage of FDM is that since no matrices are formed hence the required processing and storage capacity of the computer is small.The solution without matrices also allows for analysis of large displacements without significant additional computer effort. 6.2.3 Boundary Element Method (BEM) In this method the rock mass is taken to be a continuum and infinitesimal deformation is usually assumed. In Boundary element method (BEM) discretisation of the problem domain is necessary for the excavation boundary only.The amount of data required to describe the problem is greatly

reduced when compared to the finite element method where the entire domain has to be discretised. The influence of the infinite rockmass is automatically considered in the analysis. The boundary element method is very efficient when the defined boundaries are of greatest concern. The numerical calculation is confined to these boundary elements. The medium inside those boundaries is typically described and simulated by partial differential equations.These equations are generally linear. BEM is most efficient for homogeneous isotropic linear elastic problems. Joints are modelled explicitly in this method . Numerical convergence is often found to be a problem for models in which the rockmass is heavily jointed.Therefore problems requiring explicit consideration of several joints are often better handled by other numerical methods . A representative program for BEM is PHASES. 6.2.4 Discrete Element Method (DEM) The Discrete Element Method is also referred to as Distinct Element Method. Contrary to FEM, FDM & BEM , the ground mass, in this method, is not modeled as a continuum. Rather, the ground mass is modeled by individual blocks that are rigid in themselves.This method is applicable if the joint displacements are much higher than the internal block deformation.In this case, the deformation of the ground mass is governed by the movement along the joints between rigid blocks.Individual blocks are free to rotate and translate . The joints are modeled explicitly. The method is based on explicit solution technique. In the explicit scheme , the approximate solution for displacement (ui) at time (t+1)is carried out in terms of known values of (ui) at the previous time level (t). The explicit solution is rather straight forward, permits step by step evaluation of (ui) directly, and does not require solution of equations.Typical available programs for this method are UDEC (Universal Distinct Element Code) and 3DEC( three dimensional version of UDEC). 6.2.5 Beam Element Method with Elastic support This method is also referred to as Coefficient of Subgrade Reaction Method.It is specially suitable to simulate lining in underground caverns.The lining is simulated by beam elements .The surrounding ground,that provides the embedment of the lining,is simulated by spring elements. Spring elements are typically oriented perpendicular to the lining,simulating the normal stresses induced to the ground from outward lining deflection. This method can be used to analyse a tunnel lining.The required computer processing and storage capacity is quite small. However the model used for beam element method with elastic support can simulate simple or very simplified ground conditions. 6.2.6 Hybrid Methods Each numerical method may be used most efficiently if combined with other numerical models. By couping individual numerical methods , the strengths of each method can be preserved while its weakness may be eliminated.The combination of individual methods and their associated

models can create a model that best describes the specific problem.As an example the continuum model can be combined with the discontinuum model. The problem domain can be divided into two areas.The far off area , away from the excavtion can be modeled as a continuum. The near field, close to the excavation can be modeled with discrete elements.This reflects the anticipated ground displacement if jointed rock is encountered and movements are not restrained by support. Since the far field area is of less concern to the engineer and the ground mass is more confined , a continuum model is justified. Another example would be the combination of boundary element with finite element method. The purpose of surrounding the finite element mesh with boundary elements is to eliminate the need for arbitrary and rigid boundary conditions.Hence , the size of finite element mesh can be reduced appreciabely.

6.3 Selection of The Most Suitable Method Each numerical method has its advantages and disadvantages. The suitability and applicability of a numerical method must be ascertained for each individual case and on the objective of the study. If the rock mass is sparsely jointed with relatively big sizes of excavation then the deformation of rockmass would be continuous and then continuum approach will be suitable.The finite element method or the finite difference method may be suitable in such a case.Whenever the deformation of rockmass is more than the deformation along the joints , the continuum approach will be suitable. However if the average spacing of the joints is of similar order as the size of the excavation then the deformation along the joints will be much higher than the internal deformation of the block. In such cases the discontinuum approach will be more suitable and distinct element method will be very advantageous.

CHAPTER 7

FAST LAGRANGIAN ANALYSIS OF CONTINUA IN 3 DIMENSIONS

(FLAC3D) ___________________________________________________________

7.1 Introduction FLAC3D is a three–dimensional explicit finite–difference program for engineering mechanics computation. It is capable of simulating the behaviour of three dimensional structures built of soil, rock or other materials that undergo plastic flow when their yield limits are reached. Materials are represented by polyhedral elements within a three-dimensional grid that is adjusted by the user to fit the shape of the object to be modeled. Each element behaves according to a prescribed linear or non-linear stress/strain law in response to the applied forces or boundary restraints. The material can yield and flow , and the grid can deform ( in large strain mode) and move with the material that is represented. The explicit , Lagrangian, calculation scheme and the mixed-discretization zoning technique used in FLAC3D ensure that plastic collapse and flow are modeled very accurately.Since no matrices are formed, large three-dimensional calculations can be made without excessive memory requirements. The drawback of the explicit formulation ( i.e., small timestep limitation and the question of required damping) are overcome by automatic inertia scaling and automatic damping that does not influence the mode of failure. FLAC3D offers an ideal analysis tool for solution of three-dimensional problems in geotechnical engineering. ( FLAC3D USER’S MANUAL, VOLUME-I)

7.2 Theoretical Background 7.2.1 Formulation of 3D Explicit Finite Difference Model FLAC3D is an explicit finite difference program to study numerically the behaviour of a continuous three dimensional medium as it reaches equilibrium or steady plastic flow. The response observed derives from a particular mathematical model on one hand and from a specific numerical implementation on the other. The mechanics of the medium are derived from general principles (defination of strains , laws of motion),and the use of constitutive equations defining the idealised material. The resulting mathematical expression is a set of partial differential equations, relating mechanical (stress) and kinematic (strain rate, velocity) variables, which are to be solved for particular geometries and properties, given specific boundary and initial conditions. An important aspect of the model is the inclusion of the equations of motion.

The method of solution in FLAC3D is characterized by the following three approaches :

1) Finite difference approach (First-order space and time derivatives of a variable are approximated by finite differences assuming linear variations of the variable over finite space and time intervals, respectively.);

2) Discrete-model approach (The continuous medium is replaced by a discrete equivalent

one in which all forces involved are concentrated at the nodes of a three dimensional mesh used in the medium representation.)

3) Dynamic-solution approach (The inertial terms in the equations of motion are used as

numerical means to reach the equilibrium state of the system under consideration.)

The laws of motion for the continuum are, by means of these approaches, transformed into discrete forms of Newton’s law at the nodes. The resulting system of ordinary differential equations is then solved numerically using an explicit finite difference approach in time. 7.2.2 Grid Discretization Among three-dimensional constant strain-rate elements, tetrahedra have the advantage of not generating hourglass deformations (i.e.,deformation patterns created by combinations of nodal velocities producing no strain rate and, thus, no nodal force increments). However, when used in the framework of plasticity, these elements do not provide for enough modes of deformation. In particular situations , for example, they cannot deform individually without change of volume as required by certain important constitutive laws.To overcome this problem a process of mixed discretization is followed in FLAC3D. The principle of the mixed discretization technique is to give the element more volumetric flexibility by proper adjustment of the first invariant of the tetrahedra strain-rate tensor.A coarser discretization in zones is superposed to the tetrahedral discretization, and the first strain rate invariant of a particular tetrahedron in a zone is evaluated as the volumetric-average value over all tetrahedra in the zone. The application of the mixed discretization process allows each individual tetrahedron to reflect the property of the zone, (volume of assembly of tetrahedra remaining constant), hence reconciling its behaviour with that predicted by the theory. 7.2.3 Numerical Implemetation The general discretization of the body into zones is performed by the user.Each zone is discretized automatically by the code into sets of tetrahedra. Then boundary conditions have to be defined by the user. The boundary conditions of the problem consist of surface tractions, concentrated loads and displacements.In addition, body forces may be given and initial stress conditions imposed.For implementation in the code , all stresses and nodal velocities are initially set to zero. Then initial stresses are applied.

The main calculation steps at each timestep are given below : 1 New strain rates are derived from nodal velocities. 2 Constitutive equations are used to calculate new stresses from the strain rates and stresses at

the previous time. 3 The equations of motion are invoked to derive new nodal velocities and displacements from

stresses and forces. The sequence is repeated at every timestep, and the maximum out of balance force in the model is monitored. This force will either approach zero, indicating that the system is reaching an equilibrium state, or it will approach a constant, non-zero value, indicating that a portion of the system is at steady-state (plastic) flow of material.

7.3 Material Models There are ten built-in material models in FLAC 3D : (1) null; (2) elastic, isotropic; (3) elastic, orthotropic; (4) elastic, transversely isotropic; (5) Drucker-Prager plasticity; (6) Mohr-Coulomb plasticity; (7) strain-hardening / softening Mohr-Coulomb plasticity; (8) ubiquitous-joint plasticity; (9) bilinear strain-hardening / softening ubiquitous-joint plasticity; and (10) modified Cam-clay plasticity. Each model is developed to represent a specific type of constitutive behavior commonly associated with geologic materials. The null model is used to represent material that is removed from the model but with the associated zones left in place. The elastic, isotropic model is valid for homogeneous, isotropic, continuous materials that exhibit linear stress-strain behavior.

The elastic, orthotropic model and the elastic, transversely isotropic model are appropriate for elastic materials that exhibit well-defined elastic anisotropy. The Drucker-Prager plasticity model is a simple failure criterion in which the shear yield stress is a function of isotropic stress. The Mohr-Coulomb plasticity model is used for materials that yield when subjected to shear loading, but the yield stress depends on the major and minor principal stresses only; the intermediate principal stress has no effect on yield. The strain-softening Mohr-Coulomb model is based upon the Mohr-Coulomb model, but is appropriate for materials that show a degradation or increase in shear strength when loaded beyond the initial failure limit. The ubiquitous-joint model applies for a Mohr-Coulomb material that exhibits a well-defined strength anisotropy. The bilinear strain-softening ubiquitous-joint model combines the strain-softening Mohr-Coulomb model with the ubiquitous-joint model. This model includes a bilinear failure envelope for both the matrix and the ubiquitous joints. The modified Cam-clay model accounts for the influence of volume change on deformability and on resistance to failure. The material models in FLAC3D

are primarily intended for applications related to geotechnical engineering —e.g., underground construction, mining, slope stability, foundations, earth and rock-fill dams. The Mohr-Coulomb model is the most applicable for general engineering studies. Also, Mohr-Coulomb parameters for cohesion and friction angle are usually available more often than other properties for geo-engineering materials. Table 5.1 presents a summary of the material models in FLAC3D

Table 7.1 Constitutive Material Model in FLAC3D (FLAC3D USER MANUAL,) Model Reresentative Material Example Application

Null void holes, excavations, regions in which material will be added at later stage

Elastic homogeneous, isotropic continuum;linear stress strain behaviour

manufactured materials (e.g.,steel) loaded below strength limit; factor of safety calculation

Orthotropic Elastic materials with three mutually perpendicular planes of elastic symmetry

columnar basalt loaded below strength limit

Transversly isotropic Elastic

thinly laminated material exhibiting elastic anisotropy (e.g.,slate)

laminated material loaded strength limit

Drucker- Prager plasticity limited applications; soft clays with low friction

common model for comparison to implicit finite-element programs

Mohr-Coulomb Plasticity

loose and cemented granular materials;soils, rock, concrete

general soil or rock mechanics (e.g., slope stabilty and underground excavation)

Strain-hardening/ Softening Mohr-Coulomb

granular materials that exhibits non-linear material hardening or softening

studies in post-failure (e.g., progressive collapse, yielding pillar, caving)

Ubiquitous-Joint thinly laminated material exhibiting strength anisotropy (e.g., slate)

excavation in closely bedded strata

Bilinear strain-hardening/softening Ubiquitous- joint

laminated materials that exhibit non-linear material hardening or softening

studies in post-failure of laminated materials

materials for which deformibility and shear strength are a function of volume change

geotechnical construction on clay

When selecting a constitutive model for a particular analysis, the following two considerations should be kept in mind: 1) the known characteristics of material being modelled; 2) the intended application of model analysis. For example, the Mohr-Coulomb model should be used when stress levels are such that failure of intact material is expected.

7.4 Problem Solving with FLAC3D In order to set up a model to run a simulation with FLAC3D, three fundamental components of a problem must be specified: a finite difference grid; constitutive behaviour and material properties; and boundary and initial conditions The grid defines the geometry of the problem. The constitutive behaviour and associated material properties dictate the type of response the model will display upon disturbance (e.g., deformational response due to excavation ). Boundary and initial conditions define the in-situ state (i.e., the condition before a change or disturbance in the problem state is introduced). After these conditions are defined in FLAC3D, the initial equilibrium-state is calculated for the model. An alteration is then made (e.g., excavate material or change boundary conditions), and the resulting response of the model is calculated. The solution is reached after a series of computational steps. The general solution procedure is shown in Fig. 7.1 7.4.1 The Finite Difference Grid The finite difference grid spans the physical domain being analyzed. The smallest possible grid that can be analyzed with FLAC3D consists of only one zone. Most problems, however, are defined by grids that consists of thousands of zones or elements.Grid generation involves adjusting and shaping the mesh to fit the shape of the physical domain.It must be remembered that finer elements give better results , however, increasing the number of elements means more computing capacity. The mesh must be created in such a way that there are fine elements in regions of high stress/strain gradients (e.g., excavation boundary). 7.4.2 Material Properties The material properties required in FLAC3D are generally categorized in two groups: i) elastic deformibility properties; and ii) strength properties. The selection of properties is often the most difficult element in the generation of a model because of a high uncertainity in the property database.The problem will always involve a data-limited system. All material models in FLAC3D, except for the transversly isotropic elastic and orthotropic elastic models, assume an isotropic material behaviour in the elastic range described by two elastic constants, bulk modulus (K) and shear modulus (G). The shear and the bulk modulus are used instead of Young’s modulus and Poisson’s ratio because it is beleived that K and G corresponds to fundamental aspects of material behaviour. 7.4.3 Boundary and Initial Conditions

The boundary conditions in a numerical model consist of the values of the field variables that are prescribed at the boundary of the numerical grid.Boundaries can be either real or artificial--- real boundaries exist in the physical object being modeled, whereas artificial boundaries are introduced to enclose the chosen number of zones. The model boundaries must be far enough away from the region of study so that model response is not adversly affected. The appropriate distance depends on the purpose of the analysis. Generally boundaries should be 3 to 5 times the size of the cavern (span or height whichever is greater ). If a displacement boundary is chosen then both stress and displacements will be underestimated while a stress boundary causes both stress and displacements to be overestimated.

7.5 Comparison with other models Both, the common method of finite element and the FLAC3D translate a set of differential equations into matrix equations for each element, relating forces at nodes to displacements at nodes.However, FLAC3D differs in the following respects: 1) The mixed discretization scheme (Marti and Cundall, 1982) is used for accurate modeling of

plastic collapse loads and plastic flows. This scheme is beleived to be physically more justifiable than the reduced integration scheme commonly used with finite elements.

2) The full dynamic equations of motion are used, even with modeling systems that are

essentially static. 3) An explicit solution scheme is used.Explicit schemes can follow arbitrary non-linearity in

stress/strain laws in almost the same computer time as linear laws.Also, since it is not necessary to store matrices so a large number of elements can be modeled with a modest memory requirement and large-strain simulation is hardly more time consuming than a small –strain run,because there is no stiffness matrice to be updated.

7.6 Limitations of FLAC3D There are two main disadvantages. They are : 1) Linear simulations run slower with FLAC3D than with equivalent finite element programs; FLAC3D is most effective when applied to non linear or large-strain problems. 2) The solution time with FLAC3D is determined by the ratio of the longest natural period to the

shortest natural period in the system being modeled. Thus models containing large disparities in elastic moduli or element sizes are very inefficient to FLAC3D.

This chapter has been compiled from the User’s Manual of FLAC 3D.

R e su ltU n sa t is fa c to ry

S te p to so lu t io n

M o re te s tsn e e d e d

Y e s

N o

S te p to E q u i l ib r iu m s ta te

M o d e l S e tu p1 . G e n e ra te g r id , d e fo rm to d e s ire d sh a p e2 . D e f in e C o n s t i tu t iv e b e h a v io u r & m a te r ia l p ro p e r t ie s .3 . S p e c i f y b o u n d a ry & in i t ia l c o n d i t io n s

P e rfo rm A lte ra t io n sfo r e x a m p le-- E x c a v a te M a te r ia l- - C h a n g e b o u n d a ry c o n d it io n s

S te p to so lu t io n

E x a m in e m o d e l re sp o n se

P a ra m e te r s tu d y n e e d e d

E n d

E x a m in e m o d e l re sp o n se

S ta r t

Fig. 7.1 General Solution Procedure

CHAPTER 8

DESCRIPTION OF THE NUMERICAL MODEL __________________________________________________________________

8.1 Modelling Guidelines Numerical models in the field of Geotechnical Engineering are essentially different from the structural models. The difference lies in the availability of data regarding the material properties. There will always be the problem of lack of data in a model in the field of rock engineering.A structural model of steel or concrete can be very accurately defined and very accurate material properties can be assigned. However it is almost impossible to know enough about rockmass to model it accurately. Hence a different approach should be adopted.( Starfield, A.M. and P.A. Cundall 1988) Always start from a simple model and the objectives of the study should be well defined.Try to identify the area of interest and the likely modes of failures which are to be studied.Make suitable assumptions to simplify the model and remember them while interpreting the results. Do not build a complicated model to start with and get confused by its results. Try to understand the mechanism by taking different runs after modifying the model.For example since the rockmass properties are difficult to ascertain, a parametric study can be carried out by changing the rockmass properties and noting the results. This will give insight as to how the material properties are going to affect the results.

Basic principle studies should be done by keeping the same geometry of the cavern and with homogeneous material properties of the surrounding rock-mass. These studies should aim at the understanding of engineering principles that determine the design requirements. For example the analysis and understanding of the stress flow in piers between caverns may help to modify the cavern layout. Since very accurate modelling of rockmass is almost impossible, it is better to simplify by introducing the major joint sets and faults and ignoring the rest for modeling. However while interpreting the results and making recommendations, all the geological facts which could not be incorporated must be considered. After getting a better understanding from the simple model, more complex models may be run to explore those neglected aspects of geology. It will then be possible to give better interpretation of the results of the complex models. Finally, back analysis must be carried out to validate the model. The actual deformations as measured by instruments must be compared with those predicted by the model. The input parameters can be refined by carrying out the back analysis. It will also be very useful if an analysis has to be carried out in future as it will provide very reliable input parameters.

In this study basically four models have been generated and simulated. These are shown in Table 8.1.

Table 8.1 Models studied

Model Number Model Description Material Model 1

Powerhouse cavern, Transformer cavern & 3 gate shafts excavated.

Linear-elastic model.

2

Powerhouse cavern, Transformer cavern & 3 gate shafts along with 3 bus tunnels and 3 draft tubes excavated.

Linear-elastic model.

3

Powerhouse cavern, Transformer cavern & 3 gate shafts excavated.

Mohr-Coulomb Plastic Model.

4

Powerhouse cavern, Transformer cavern & 3 gate shafts along with 3 bus tunnels and 3 draft tubes excavated.

Mohr-Coulomb Plastic Model.

The machine hall and the transformer cavern have been excavated in four stages and the gate shaft in one stage. All the bus tunnels has been excavated in the third stage and the draft tubes has been excavated in the third stage (in models 2 & 4). The order of excavations are:

1 Excavation of powerhouse & transformer cavern ; 2 Excavation of machine hall bench-1 (EL 380.5-367.3m), transformer cavern bench-1

(EL. 397.5-385.8m )

3 Excavation of machine hall bench 2 (EL. 367.3-352m), transformer cavern bench 2 ( EL. 385.8-376.5m ), & 3 gate shafts bench ( EL. 376.5-367.3m). Excavation of all 3 bus tunnels (in models 2 & 4 ).

4 Excavation of machine hall powerhouse bench 3 (EL. 352-340m), transformer

chamber bench 3 (EL: 376.5-374.5m). The flow chart of the model set-up and solution procedure is shown in Fig. 8.1.

Fig. 8.1 Flow chart for problem set-up and solution

1 Set up the boundaries of model2. Create the geometry of the caverns,draft tubes & bus tunnels3. Define the material properties and models4. Specify boundary and initial conditions

Excavation of bench 2 of both the caverns and 3 gate shafts.

Excavation of 3 bus tunnels and 3 draft tubes in models 2 & 4

Excavation of crowns of both machine hall and transformer

cavern

Excavation of bench 1 of both caverns i.e. machine hall and transformer caverns

Excavation of bench 3 of powerhouse & transformer

cavern

FINAL SOLUTION

Timestep to equilibrium





8.2 Model Geometry The ground surface elevation at the powerhouse area varies from El 550 m to 580 m, while the rock surface elevation ranges from 530m to 550m. For building up the model a flat ground surface has been assumed with ground elevation as 540m. The underground power-house machinehall is 100.5 m long, 46.5m high and 22 m in span. However in this study only the three units of the power-house have been modelled which are 72.5 m in length ( 22.5x3 +5 = 72.5m) and not the full length of 100.5 m. The service bay portion of 28m length and control room area etc. has not been included in the model. This has been done to reduce the model size and hence the computer capacity for problem solving without affecting the objectives of the study. The unit bay portion of the powerhouse, which has been modeled, is the deeper portion of 46.5m while the remaining portion is of considerably less height hence not so important from the stability point of view. The transformer cavern is running parallel to the machine hall only for the unit bay portion (72.5m). Hence, the stress flow in the piers between the caverns is valid only for the bay portion, which have been modelled. The powerhouse machine hall is connected to the transformer cavern by three bus tunnels while the powerhouse cavern is connected to the gate shaft in transformer cavern by three draft tubes. As discussed earlier the model 1 and 3 consist of only the two caverns . Models 2 and 4 have all the two caverns, the three bus tunnels and all the three draft tubes excavated. The models 1 and 3, where neither bus tunnels nor draft tubes have been excavated, are like 2-dimensional models. When the results of model 1 and 2 will be compared then the effect of excavation of the bus tunnels and the draft tubes can be studied. Similarly when models 1 & 2 are compared with models 3 & 4 then the effect of non linearity can be studied. The model geometry has been carefully created. The model is 600m wide ( 300m on either side of the longitudinal axis of the powerhouse), 473m long ( 200m on either side of the six units of the cavern) and 540m high ( 153m above the crown of the powerhouse and 340 m below the powerhouse). The transformer cavern is 72.5 m long and parallel to the powerhouse. The model geometry is the same for all four models which are studied is shown in Fig 8.1. The Fig 8.2 shows all the excavations carried out in models 2 & 4 , consisting of the two caverns, all three bus tunnels and the three draft tubes.

8.3 Boundary Conditions The boundary of the model has been chosen at a sufficient distance away from the excavation area to eliminate the boundary effect. The model boundaries are 340 m below the powerhouse, (5 times the height of the powerhouse), 200 m on either side of the longitudinal axis of the powerhouse which is much greater than 4 times the span or height of the largest cavern . In the length direction also, the boundaries are 200m away from the excavation which is much larger than 4 times the span or height of the largest cavern.

FLAC3D 2.00

SINTEF Civil and Environ. Eng.Rock and Mineral Engineering

Step 10988 Model Perspective19:15:01 Thu Jun 1 2000

Center: X: 3.000e+002 Y: 3.625e+001 Z: 2.700e+002

Rotation: X: 20.000 Y: 0.000 Z: 30.000

Dist: 2.132e+003 Mag.: 1Ang.: 22.500

View Title: Fig. 8.1 Full Model

Surface

FLAC3D 2.00


Step 12226 Model Perspective19:23:37 Thu Jun 1 2000

Center: X: 3.300e+002 Y: 3.625e+001 Z: 3.800e+002

Rotation: X: 9.851 Y: 358.272 Z: 15.000

Dist: 1.895e+003 Mag.: 4.77Ang.: 22.500

View Title: Fig 8.2 Full Excavated Portion of Model (Model 2 & 4)

Surface Null zones only

Roller displacement boundaries has been applied on all five sides leaving the surface as free. A displacement boundary causes the stress and displacement to be underestimated. This fact has to be considered while interpreting the result. The boundary condition is shown in Fig. 8.4

Fig. 8.4 Model with displacement boundaries

8.4 In-situ Stress Field According to the in-situ stress measurements, the vertical stresses are low but looking into the topography and with limited data available vertical stresses are taken as equal to the gravitational stresses and the horizontal stresses are taken equal to the vertical stresses. Hence the lateral coefficient Ko has been adopted as 1.0 for this study.The rock cover over the powerhouse is between 140 and 160 m, hence the vertical insitu stress is around 4.32 MPa and the horizontal stress is 4.32 MPa. The in-situ stresses are thus moderate.

8.5 Mechanical Properties of Rock As already discussed, the different rock layers has not been modelled in this study. This study is a continuum analysis without incorporation of discontinuities. The powerhouse, transformer cavern and the gate chamber falls in phyllitic quartzites rock.

As the data regarding the rock strength parameters was not available, hence the Elastic modulus was calculated using Barton’s formula based on Q value of rock.

. Em = 25 log 10 Q

Using the average value of Q in powerhouse area the Em value works out as 12 GPa and Shear modulus is calculated based on Poisson’s ratio.

In this study models 1 & 2 has been analysed by using a linear elastic model. For the elastic model the bulk modulus (K) and shear modulus (G) has been specified. The Bulk modulus , K , and shear modulus , G , are related to Young’s modulus, E , and Poisson’s ratio, ν, by the following equations.

)21(3 ν−= EK

)1(2 ν+= EG

GKKGE+

=39

)3(223GKGK

+−=ν

Models 3 & 4 have been analysed by using Mohr-Coulomb plasticity model. For the Mohr- Coulomb plasticity model, the required properties are :

1) bulk and shear modulii, 2) friction and dilation angles, 3) cohesion, 4) tensile strength

8.6 Material Models used in Analysis As already discussed in section 7.3 of Chapter 7 there are ten material models available in FLAC3D. However in this study analysis has been carried out using two material models. Models 1, & 2 have been analysed using the linear elastic model. Models 3 & 4 have been analysed using the Mohr- Coulomb plasticity model . The linear elastic model uses the linear reversible Hooke’s law of elasticity. The relationship between the Young’s modulus, bulk modulus, shear modulus and the Poisson’s ratio has already been stated in section 8.5 of this chapter.

The Mohr-Coulomb plasticity model represents a material that yields when subjected to shear loading. The shear yield function fs is : fs = σ1-σ3 Nϕ +2c(Nϕ)0.5 where Nϕ = (1+Sinϕ)/(1-Sinϕ), σ1 = major principal stress, σ3 = minor principal stress

ϕ = Friction Angle, c = cohesion. Shear yield is detected if fs < 0 (tension is taken as positive).Then plastic flow is allowed to occur in order to restore the condition fs = 0.

CHAPTER 9

MODELLING RESULTS, COMPARISON AND DISCUSSION

9.1 MODELLING RESULTS As already discussed in chapter 8 that four models have been analysed in this study. The geometry and the boundaries are the same for all the models. Two models have been analysed assuming elastic behaviour of rock and the other two models have been analysed using Mohr-Coulomb plasticity model. Selected relevant results simulated from these four models have been presented in this chapter while some other results are shown in the Appendices A1 to A5. 9.1.1 Results of Model 1 It is a linear elastic model in which both the machine hall cavern and transformer caverns and all three draft tube gate shafts are fully excavated. The major principal stresses at the power-house location at RD 36.25m before any excavation has been done is shown in Fig. 9.1. The RD 36.25 m lies exactly at the centre of the powerhouse cavern and hence results at this section are the best representation of the major Principal stresses, as it is not affected by the boundaries etc. The major principal stress is compressive through out the section before excavation. Fig 9.2 shows the major principal stress after both the caverns and all the three draft tube gate shafts have been excavated at RD 36.25m. The area between the machine hall cavern and transformer cavern has low stresses. It represents the stress redistribution after excavation. There is de-stressing on both the walls of the powerhouse machine hall as well as on the transformer chamber. The stress plot indicates that the major principal stress at certain local areas is tensile close to the upstream and downstream walls of the machine hall cavern from El 374.5 m to El 340m. The displacement vector at RD 36.25m is shown in Fig. 9.3. The maximum deformation is 2.164 cm at about the centre of walls of the machine hall cavern. The deformation vector indicates that the rock mass is not displaced at the machine hall cavern near the crown. This can be explained by the relatively moderate K0 value of 1.0 (σh/σv) and consequently moderate horizontal stress. Fig 9.4 shows the major Principal stresses at RD 36.25m. The compressive stresses are 18 MPa to 24 MPa near the crown of both caverns and bottom of the cavern. The stress plot indicates that the stress near the centre of caverns varies from 7 MPa to 12 MPa. The major principal stress and displacement vectors are shown in plan at El 380.50 m in Appendix A1-1 and A1-2 respectively. Elevation 380.50m has been chosen because both the caverns are excavated at EL 380.50. The Fig A1-1 indicates the Major Principal stresses at El 380.5m. It shows that that there are low compressive stresses near the centre of cavern close to both the walls of the machine hall cavern and the transformer cavern as the major principal stress varies from 9.3 Mpa to 12 Mpa.. The Appendix A1-2 shows that most of the displacement vectors at EL. 380.50 m, which is on the walls of the machine hall and transformer cavern. It indicates that the maximum displacement is simulated about 1.5 cm at EL. 380.50 m, which occurs at the centre of both the caverns.

FLAC3D 2.00


Step 7173 Model Perspective15:19:49 Thu May 11 2000

Center: X: 3.324e+002 Y: 3.355e+001 Z: 3.659e+002

Rotation: X: 357.104 Y: 0.000 Z: 0.982

Dist: 1.895e+003 Mag.: 5.8Ang.: 22.500

Plane Origin: X: 0.000e+000 Y: 3.625e+001 Z: 0.000e+000

Plane Normal: X: 0.000e+000 Y: 1.000e+000 Z: 2.833e-016

View Title: Fig 9.1 Principal stresses at RD 36.25m before Excavation (Model 1)

Grid Plane: on Linestyle

Principal Stresses Plane: on Local face system Compression Linestyle Maximum = 2.120e+007

FLAC3D 2.00


Step 12235 Model Perspective18:10:46 Wed May 17 2000

Center: X: 3.246e+002 Y: 3.447e+001 Z: 3.484e+002

Rotation: X: 357.631 Y: 0.000 Z: 0.744

Dist: 1.895e+003 Mag.: 4.02Ang.: 22.500


Plane Orientation: Dip: 90.000 DD: 0.000

View Title: Fig 9.2 Principal Stress Vectors at RD 36.25m (Model 1)

Principal Stresses Plane: on Local face system Compression Linestyle Tension

Maximum = 2.478e+007 Linestyle

FLAC3D 2.00



Center: X: 3.246e+002 Y: 3.385e+001 Z: 3.622e+002

Rotation: X: 357.215 Y: 0.000 Z: 0.744

Dist: 1.895e+003 Mag.: 5.21Ang.: 22.500



View Title: Fig 9.3 Displacement Vectors at RD 36.25 m (Model 1)

Displacement Plane: on Maximum = 2.164e-002 Linestyle


FLAC3D 2.00



Center: X: 3.246e+002 Y: 3.385e+001 Z: 3.622e+002

Rotation: X: 357.215 Y: 0.000 Z: 0.744

Dist: 1.895e+003 Mag.: 5.21Ang.: 22.500



View Title: Fig 9.4 Major Principal Stresses at RD 36.25 m (Model 1)

Contour of SMin Plane: on Gradient Calculation

-2.3461e+007 to -2.0000e+007-2.0000e+007 to -1.8000e+007-1.8000e+007 to -1.6000e+007-1.6000e+007 to -1.4000e+007-1.4000e+007 to -1.2000e+007-1.2000e+007 to -1.0000e+007-1.0000e+007 to -8.0000e+006-8.0000e+006 to -7.4366e+006

Interval = 2.0e+006


The principal stresses vectors at EL. 380.50 m and EL 363.25 m in model 1 are shown in Appendix A1-3 and Appendix A1-4. Fig A1-3 shows maximum compressive stress of 28 MPa and no tensile stresses at EL 380.50 m. Fig A1-4 shows maximum compressive stress of 34.6 MPa and low tensile stresses at some local areas at EL 363.25 m near the walls of machine hall cavern. 9.1.2 Results of Model 2 Model 2 is the linear elastic model in which both the caverns and all three bus- tunnels and three draft tubes have been excavated. The excavation has been made in four stages as already discussed in chapter 7. Fig 9.5 and 9.6 indicates the Major Principal stresses at RD. 34.5m which is section passing through the centre draft tube of middle unit and at RD 25.5m which is section passing through the centre of bus duct gallery of the second unit. It indicates high compressive principal stress of about 18 to 24 MPa near the crown of both the caverns and near the bottom of both the caverns. It also indicates low compressive stress areas of 6 to 10 MPa close to centre of cavern walls in both machine hall cavern and transformer cavern. The major principal stress at RD 34.5m is shown in Fig. 9.5. The major principal stress plot at RD 34.5m shows that the low-stress area near centre of both walls of the machine hall is in larger extent as compared to the transformer cavern.. The major principal stress at RD 25.5m (through the second bus tunnel ) is shown in Fig 9.6. It shows some low stress areas close to the bus bar area between both the caverns. Fig 9.7 shows the deformation against the steps (stages of excavation) for two points. Point 4 is at the crown of the machine hall of powerhouse while point 2 is on the downstream wall of the powerhouse cavern (El 363.20 m). This figure shows that most of the displacement at the point 4, which is on the crown of the powerhouse, occurs immediately at the first heading and thereafter almost remains constant as excavation proceeds. While the displacement at point 2, which is on the wall of the powerhouse, goes on increasing as the benching proceeds. Point 8 is at the crown of the transformer chamber, while point 6 is on the upstream wall of the transformer chamber (El 387.0 m). While the displacement at point 6, which is on the upstream wall of the transformer chamber, goes on increasing as the benching proceeds but it is considerably less than the deformation in machine hall downstream wall due to small cavern size. It also shows that most of the displacement at the point 8, which is on the crown of the transformer chamber, occurs immediately at the first heading and thereafter almost remains constant as excavation proceeds and is much smaller in magnitude for smaller span. Table 9.1 summarizes the maximum displacement at different stages. The maximum displacements at crown and walls are tabulated. All the displacement values are those occuring at RD 36.25 m ( centre of powerhouse). It is observed that the displacement of walls is much higher than the displacement at the crown. The displacement of walls continues as the excavation proceeds. Most of the displacement at crown occurs immediately in the first stage and then almost remains constant. In fact the displacement of crown at the third stage is highest.

FLAC3D 2.00



Center: X: 3.375e+002 Y: 3.344e+001 Z: 3.662e+002

Rotation: X: 357.096 Y: 0.000 Z: 1.133

Dist: 1.897e+003 Mag.: 6.18Ang.: 22.500





-2.3322e+007 to -2.0000e+007-2.0000e+007 to -1.8000e+007-1.8000e+007 to -1.6000e+007-1.6000e+007 to -1.4000e+007-1.4000e+007 to -1.2000e+007-1.2000e+007 to -1.0000e+007-1.0000e+007 to -8.0000e+006-8.0000e+006 to -6.0000e+006-6.0000e+006 to -5.6811e+006

Interval = 2.0e+006


FLAC3D 2.00



Center: X: 3.375e+002 Y: 3.344e+001 Z: 3.662e+002

Rotation: X: 357.096 Y: 0.000 Z: 1.133

Dist: 1.897e+003 Mag.: 6.18Ang.: 22.500





-2.3551e+007 to -2.0000e+007-2.0000e+007 to -1.8000e+007-1.8000e+007 to -1.6000e+007-1.6000e+007 to -1.4000e+007-1.4000e+007 to -1.2000e+007-1.2000e+007 to -1.0000e+007-1.0000e+007 to -8.0000e+006-8.0000e+006 to -6.1106e+006

Interval = 2.0e+006


FLAC3D 2.00



Center: X: 3.000e+002 Y: 3.625e+001 Z: 2.700e+002

Rotation: X: 0.000 Y: 0.000 Z: 0.000

Dist: 1.895e+003 Mag.: 1Ang.: 22.500



View Title: Fig9.5 Deformations at Various Stages of excavation

History Location

2

4 68

Boundary Plane: on Null zones only Linestyle

History

0.8 0.9 1.0 1.1 1.2

x10e4

0.2

0.4

0.6

0.8

1.0

1.2

1.4

1.6

1.8

x10e-2

2 Displacement Mag. Gp 13235 Linestyle 2.010e-006 <-> 1.876e-002

2.711e-004 <-> 3.063e-003 6 Displacement Mag. Gp 13356 Linestyle 2.090e-006 <-> 8.863e-003 8 Displacement Mag. Gp 13303

4 Displacement Mag. Gp 13111 Linestyle

FLAC3D 2.00



Center: X: 3.411e+002 Y: 3.411e+001 Z: 3.500e+002

Rotation: X: 357.583 Y: 0.000 Z: 1.244

Dist: 1.895e+003 Mag.: 5.7Ang.: 22.500



View Title: Fig9.5 Displacement vectors at RD 34.5m (Model 2)



The displacement of walls is higher because the height of the walls is greater than the span of the powerhouse. Fig. 9.8 shows the deformation at RD 34.5m in model 2 at the final stage. The maximum deformation is 2.16cm.

Table 9.1 Maximum Displacements at different stages.

Location 1st stage (cm)

2nd stage (cm)

3rd stage (cm)

4th stage (cm)

Machine hall Crown

2.50 2.00 3.00 3.50

Machine hall Walls

1.00 2.00 14.00 21.57

Transformer Chamber Crown

1.20 1.00 1.00 1.00

Transformer Chamber

Walls

0.5 2.00 8.00 8.50

The displacement vector at RD 34.50m is shown in Fig. 9.8. The maximum deformation is 2.16 cm at about the centre of walls of the machine hall cavern. The deformation vector indicates that the rock mass is not displaced at the machine hall cavern near the crown. This can be explained by the relatively moderate K0 value of 1.0 (σh/σv) and consequently moderate horizontal stress. The Major principal stresses and displacement vectors at EL. 380.50m in model 2 are shown in Appendix A2-1 and A2-2 respectively. The Fig A2-1 indicates the major principal stresses are low close to te centre of cavern and close to three bus bar plots show that due to stress redistribution. The displacement vectors in Fig. A.2-2 at EL. 380.50 m shows maximum displacement near the centre of both the caverns is about 1.44 cm. The displacement vectors and Principal stress vectors at EL 363.25m are shown in Appendix A.2-3 and Appendix A.2-4. Appendix A.2-3 shows maximum displacement near the centre upsteam wall of the machine hall cavern is about 2.16 cm.. Fig A1-4 shows maximum compressive stress of 34.9 Mpa and low tensile stresses at local areas at EL 363.25 m near the centre of walls of machine hall cavern and close to bus bar. 9.1.3 Results of Model 3 Model 3 is the one in which all the three caverns are excavated and the rockmass is assumed to behave as per the Mohr- Coulomb’s plastic material. The major principal stress at RD 36.25m is shown in Fig 9.9. Stress redistribution after the excavation of the both caverns shows low stresses near walls and high streses above the crown. The major principal stress at some local areas is tensile near the centre of machine hall cavern walls and it varies between 2 Mpa to 0 Mpa on both the walls. However near the crown the major principal stress varies between 2.5 MPa to 10 MPa. The displacement vectors at RD 36.25 m is shown in Fig. 9.10. The maximum displacements are shown as 4.28 cm near the centre of the machine hall wall.

FLAC3D 2.10Step 11794 Model Perspective12:01:48 Tue Aug 06 2002

Center: X: 3.332e+002 Y: 3.373e+001 Z: 3.620e+002

Rotation: X: 357.223 Y: 0.000 Z: 1.005

Dist: 1.896e+003 Mag.: 5.01Ang.: 22.500



View Title: Fig. 9.9 Major Principal Stresses at RD 36.25 m ( Model 3)

Contour of SMin Plane: on Magfac = 0.000e+000 Gradient Calculation

-1.8728e+007 to -1.5000e+007-1.5000e+007 to -1.2500e+007-1.2500e+007 to -1.0000e+007-1.0000e+007 to -7.5000e+006-7.5000e+006 to -5.0000e+006-5.0000e+006 to -2.5000e+006-2.5000e+006 to 0.0000e+000 0.0000e+000 to 1.8214e+006

Interval = 2.5e+006

Grid Plane: on

FLAC3D 2.10Step 11794 Model Perspective11:54:31 Tue Aug 06 2002

Center: X: 3.332e+002 Y: 3.373e+001 Z: 3.620e+002

Rotation: X: 357.223 Y: 0.000 Z: 1.005

Dist: 1.896e+003 Mag.: 5.01Ang.: 22.500



View Title: Fig. 9.10 Displacement Vectors at RD 36.25 m ( Model 3)

Grid Plane: on Magfac = 5.500e+000 Exaggerated Grid Distortion Linestyle


FLAC3D 2.00



Center: X: 3.319e+002 Y: 3.402e+001 Z: 3.563e+002

Rotation: X: 357.396 Y: 359.990 Z: 0.964

Dist: 1.897e+003 Mag.: 5.55Ang.: 22.500



View Title: Fig 9.11 Principal Stress Vectors at RD 36.25m (Model 3)




6 Displacement Mag. Gp 13356 Linestyle

FLAC3D 2.00



Center: X: 3.319e+002 Y: 3.402e+001 Z: 3.563e+002

Rotation: X: 357.396 Y: 359.990 Z: 0.964

Dist: 1.897e+003 Mag.: 5.55Ang.: 22.500



View Title: Fig 9.12 Deformations at Various stages at RD 36.25m (Model 3)

History

0.7 0.8 0.9 1.0 1.1

x10e4

0.2

0.4

0.6

0.8

1.0

1.2

1.4

1.6

1.8

2.0

2.2

2.4

2.6

2.8

3.0

3.2

3.4

3.6

3.8x10e-2

2 Displacement Mag. Gp 13235 Linestyle 5.638e-008 <-> 3.810e-002

2.465e-008 <-> 2.010e-002 8 Displacement Mag. Gp 13303 Linestyle 2.354e-005 <-> 8.477e-003 4 Displacement Mag. Gp 13111 Linestyle 2.614e-005 <-> 1.763e-002

Vs. Step 6.050e+003 <-> 1.190e+004

2

4 6

8

Fig 9.11 shows the principal stress vectors at RD 36.25m for model 3. It shows that the maximum compressive stress is 25.8Mpa and is tensile at certain local areas close to the centre of cavern. The Fig. 9.12 indicates failure state of the model. The plot shows locations on the downstream and the upstream wall of the powerhouse, where permissible shear stress has been exceeded in the past ( during some stage of excavation) but are stable in the final stage. Some shear failure in the final stage is shown in the floor and lower portion of the downstream wall.

The major principal stresses at EL 380.50m in model 3 are shown in Appendix A 3-1.This plot shows both the major principal stresses. The major principal stress is about 3 Mpa to 6 Mpa at the centre of caverns. Fig A 3-2 shows the displacement vectors at El 380.50m. It has the maximum displacement of about 2.98 cm near the centre of cavern at EL. 380.50m Fig A.3-2 shows the principal stresses at El 380.50m and its magnitude at various locations. This plot indicates the stress state in the pillars between the machine hall of powerhouse and the transformer chamber and also at the walls of the powerhouse and the transformer chamber. Figure A. 3-4 indicates the failure state during the construction stage due to shear forces. 9.1.4 Results of Model 4 Model 4 is the one in which all the three caverns and all six draft tubes and bus tunnels are excavated and the rockmass is assumed to behave as per Mohr- Coulomb plasticity model. The major principal stress at RD 34.50m is shown in Fig. 9.13. It indicates the stress redistribution after the full excavation. The destressed zone is large on the downstream side of the machine hall of powerhouse. The major principal stress at some local areas is tensile on the centre of walls of the powerhouse from El 370 m to El 349 m Fig 9.14 shows the displacement at RD 34.50m and it has the maximum displacement of 4.3 cm near the centre of machine hall cavern. Fig. 9.15 indicates the principal stress vectors at RD 34.50 m. It shows maximum compressive stress of 25.8 Mpa and tensile stresses at some local areas close to the crown. The failure states are shown at Ch 68m (through the bus tunnel & draft tube) in model 4 in Fig 7.16. It shows all those elements which may fail in shear and in tension. This shows that some elements exceeded their permissible shear stress in past on the walls of both the caverns but on the crown. Similarly much portion of the draft tube roof and floor seems to have exceeded the permissible shear stress in past. At certain locations on the draft tube roof as shown in the Fig. 9.16, the shear stress has been exceeded in past as well as in the present stage also. The Major principal stresses and displacement at El 380.50 m are shown in model 4 in Appendix A4-1 and A4-2 respectively. The principal stress and failure block state are shown at a plan at El 363.25m in Appendix A4-3 and Appendix A.4-4.

FLAC3D 2.00



Center: X: 3.365e+002 Y: 3.349e+001 Z: 3.655e+002

Rotation: X: 357.118 Y: 0.026 Z: 1.105

Dist: 1.895e+003 Mag.: 6.35Ang.: 22.500



View Title: Fig 9.13 Major Principal Stresses at RD 34.5m (Model 4)


-1.8415e+007 to -1.5000e+007-1.5000e+007 to -1.2500e+007-1.2500e+007 to -1.0000e+007-1.0000e+007 to -7.5000e+006-7.5000e+006 to -5.0000e+006-5.0000e+006 to -2.5000e+006-2.5000e+006 to 0.0000e+000 0.0000e+000 to 1.6008e+006

Interval = 2.5e+006


FLAC3D 2.00



Center: X: 3.365e+002 Y: 3.349e+001 Z: 3.655e+002

Rotation: X: 357.118 Y: 0.026 Z: 1.105

Dist: 1.895e+003 Mag.: 6.35Ang.: 22.500



View Title: Fig 9.14 Displacement Vectors at RD 34.5m (Model 4)



FLAC3D 2.00



Center: X: 3.365e+002 Y: 3.349e+001 Z: 3.655e+002

Rotation: X: 357.118 Y: 0.026 Z: 1.105

Dist: 1.895e+003 Mag.: 6.35Ang.: 22.500



View Title: Fig 9.15 Principal Stresses vectors at RD 34.5m (Model 4)




9.2 FLAC3D 2.00



Center: X: 3.365e+002 Y: 3.349e+001 Z: 3.655e+002

Rotation: X: 357.118 Y: 0.026 Z: 1.105

Dist: 1.895e+003 Mag.: 6.35Ang.: 22.500



View Title: Fig 9.16 Failure State at RD 34.5m (Model 4)


Block State Plane: on

Noneshear-n shear-pshear-n shear-p tension-pshear-p

COMPARISON OF RESULTS OF NUMERICAL MODELS The various model results simulated are compared and analysed to evaluate the results obtained. 9.2.1 Comparison of the Elastic Models (Model 1 & 2) Models 1 and 2 have been simulated using FLAC3D as Elastic models. The main difference between models 1 & 2 is that in model 2 the excavation includes the full excavation of all the three bus tunnels and all the three draft tubes. The comparison of the results of these two models indicates the effect of excavation of the bus tunnels and the draft tubes on the stability of the three caverns. The major principal stresses are compared at the same locations at RD 36.25m in Model 1 and Model 2 and it is found that low stress area is larger in extent in model 2 in which the draft tubes and the bus tunnels are also excavated. The increase in the de-stressed region is on the downstream side of the machine hall cavern between machine hall cavern and the transformer chamber and also close to the floor of the transformer chamber. However the stress condition is almost similar in both the models on the upstream side of the powerhouse cavern. This indicates the effect of draft tube and bus tunnel excavation on the stress situation. The maximum deformations at the crown and walls in both models are compared in Table 9.2.

Table 9.2 Maximum Deformation at Crown and Walls (Models 1 & 2) Location Maximum deformation

in model 1 (cm)

Maximum deformation in model 2

(cm) Crown 0.31 0.31 Machine hall

Cavern Wall 1.87 1.88 Crown 0.13 0.13 Transformer

chamber Wall 0.89 0.89

There is a slight change in deformation due to the additional simulation of excavation of bus tunnels and the draft tubes in Model 2 in addition to both caverns and all the three draft tube gate shafts for Model 1. The deformation of walls is much greater than that of crown for the machine hall as well as transformer chamber as the height of both these caverns are greater than their respective spans. However, in transformer chamber as the ratio of height and span is less than then ratio of height and span for the machine hall cavern and hence wall deformations in transformer chamber are much less than the wall deformations for walls of the machine hall cavern. It is interesting to note that due to the excavation of draft tube and bus tunnels the change in wall deformation is more than in crown deformation. The excavation of draft tubes and bus tunnels lead to deformations, which may more affect the walls as compared to the crown. The total deformation of walls is much higher than deformation of crown. Hence, the stability of walls will require more attention than crown.

9.2.2 Comparison of the Plastic Models (Models 3 & 4) On comparing the major principal stress at the same location, i.e. RD 36.25 m, it is found that de-stressing is much larger in extent, in model 4, in which the draft tubes and the bus tunnels are also excavated. The increase in the de-stressed region is on the downstream side of the powerhouse cavern between the machine hall cavern and the transformer chamber and also on the floor of the transformer chamber. However the stress condition is almost similar in both the models on the upstream side of the machine hall cavern. This indicates the effect of draft tube and bus tunnel excavation on the stress situation. The maximum deformations at the crown and walls in both models are compared in Table 9.3.

Table 9.3 Maximum Deformation at Crown & Walls (Model 3 & 4) Location Maximum Deformation

in Model 3 (cm)

Maximum deformation in Model 4

(cm) Crown 1.76 1.81 Powerhouse


chamber Wall 2.01 2.25 There is an increase in deformations due to the excavation of bus tunnels and the draft tubes. It can be noticed that in transformer chamber & gate chamber, the increase in wall deformation is more than crown deformation due to excavation of bus tunnels and draft tubes. This is due to the fact that draft tube and bus tunnel excavation induces deformations. However, in the machine hall cavern, the deformation in both crown as well as wall increases almost in same ratio. This can be explained by the fact that many elements near the walls have already exceeded their permissible shear and tensile stress in the past and hence the excavation of bus tunnels and draft tubes leads to increase in deformation in walls of the powerhouse. 9.2.3 Results of the Elastic & Plastic Models (Model 1 & 3) The Model 1 has been simulated as Elastic model and model 3 has been simulated as Mohr-Coulomb’s plastic model. The difference between models 1 & 3 is the different material models simulated. The comparison of the results of these two models indicates the effect of non linear behaviour of rock mass. On comparing the major principal stress at RD 36.25m in both the models, it has been found that the de-stressed zone is slightly more in the plastic model (Model 3). The magnitude of the tensile stresses is also slightly higher in the plastic model. This is due to the effect of non-linearity. In the plastic model, once the stress in the rock-mass reaches the yield value it no longer takes any load and any further stress is distributed to the adjoining rock-mass. Thus the de-stressed zone in the plastic model is more than in the elastic model. Fig 9.12 shows the plastic states and the elements, which have exceeded the permissible shear or tensile stress in past or in present state. Thus the stress redistribution will be different in plastic model as

compared to the elastic model. The plastic model is much closer to the natural rock-mass model provided the values of material properties simulated represents the actual material properties of the rock-mass. The maximum deformations at the crown and walls in both models are compared in Table 9.4.

Table 9.4 Maximum Deformation at Crown and Walls (Models 1 & 3) Location Maximum Deformation

in Model 1 (cm)

Maximum Deformation in Model 3



chamber Wall 0.89 2.01 The deformation in Plastic model is higher than Elastic model. The deformation in crown in plastic model is much higher than deformation in walls. It is due to plastic properties of material. This can be explained by the fact that many elements near the walls have already exceeded their permissible shear and tensile stress in the past and hence the excavation of cavern in stages leads to increase in deformation in walls of the powerhouse. 9.2.4 Results of the Elastic & Plastic Models (Model 2 & 4) The model 2 has been simulated as Elastic model and model 4 has been simulated as Mohr-Coulomb’s plastic model. The difference between models 2 & 4 is the different material models simulated. The comparison of the results of these two models indicates the effect of non linear behaviour of rock mass. On comparing the major principal stress at RD 36.25m in both the models, it has been found that the de-stressed zone is slightly more in the plastic model (Model 4). The magnitude of the tensile stresses is also slightly higher in the plastic model. This is due to the effect of non-linearity. In the plastic model, once the stress in the rock-mass reaches the yield value it no longer takes any load and any further stress is distributed to the adjoining rock-mass. Thus the de-stressed zone in the plastic model is more than in the elastic model. Fig 9.16 shows the plastic states and the elements, which have exceeded the permissible shear or tensile stress in past or in present state. Thus the stress redistribution will be different in plastic model as compared to the elastic model. The plastic model is much closer to the natural rock-mass model provided the values of material properties simulated represents the actual material properties of the rock-mass. The maximum deformations at the crown and walls in both models are compared in Table 9.5

Table 9.5 Maximum Deformation at Crown and Walls (Models 2 & 4) Location Maximum Deformation

in Model 2 (cm)

Maximum Deformation in Model 4



chamber Wall 0.89 2.25

There is an increase in deformations due to the excavation of bus tunnels and the draft tubes. It can be noticed that in transformer chamber & gate chamber, the increase in wall deformation is more than crown deformation due to excavation of bus tunnels and draft tubes. This is due to the fact that draft tube and bus tunnel excavation induces deformations. However, in the machine hall cavern, the deformation in both crown as well as wall increases almost in same ratio. This can be explained by the fact that many elements near the walls have already exceeded their permissible shear and tensile stress in the past and hence the excavation of bus tunnels and draft tubes leads to increase in deformation in walls of the powerhouse.

9.3 Discussion about Rock Support Rock supports are inevitable in large caverns. The self-supporting capacity of the rock-mass should be used to its advantage, and the amount of rock support should be kept at minimum. The approach to design the rock support system is one of the following: (Nilsen, B. & Thideman, A., 1993)

1) Empirical Method 2) Classification systems 3) Analytical methods

In Teesta-V Hydro-electric project, a permanent support system of shotcreting and rock bolting has been proposed for the power-house and the transformer chamber. A permanent support system of concrete lining has been adopted for the bus tunnels, tail race tunnels and all tunnel intersections. The details of support system provided in the powerhouse and the transformer chamber are shown in Table 5.5 and 5.6 in Chapter 5. Shotcreting and rock bolting support system has been designed according to the New Austrian tunnelling Method (NATM), a construction plan which follows the sequence of “Design-Construction-Monitoring-Revised Design”. The Q-system has been proposed to evaluate the rock support and has been checked with the empirical methods and numerical methods. According to Q-system, the Q value of rock for most areas of the power house cavern is fair to good (3-38) and the excavation support ratio of 1 was adopted for powerhouse. Based on the Q- method, the rock support for powerhouse is as below:

1. Fair to Good rock area- Support with Spot rock-bolts- 6 m long, 25 mm in diameter in about 70 % length of cavern

2. Poor to Fair rock: Systematic rock-bolts- 6 m long, 25 mm in diameter @ 2m X 2 m along with 50-90 mm thick shotcrete in about 20% length of cavern

3. Poor to very poor rock: 100-150 mm thick Fibre reinforced shotcrete and spotbolting is proposed.

There are many empirical methods to determine rock supports and this is the most commonly used method. Other methods are mostly applied as a supplement to this method. The U.S. Army (1980) suggests empirical methods of assigning length and spacing of rock bolts which is shown in Table 9.6. Table 9.6 Minimum bolt length and maximum spacing for rock reinforcement (Sinha, R.,S., etc. 1980)

a Two times the bolt spacing b

Three times the width of critical and potentially unstable rock blocks.

For elements above springing line 1) Spans less than 6.6 m---1/2 span 2) Spans from 20m to 330 m—1/4 span

c

3) Spans 6.6 m to 20 m--- interpolate between 3.3m and 5m. For elements below springing line 1) For openings less than 20m high use lengths as determined in c above

Minimum Bolt Length

Greatest of

d

2) For openings greater than 20 m – 1/5 the height. a ½ the bolt length b 1—1/2 the width of critical and potentially unstable rock

blocks.

Maximum Spacing

Least of

c 2m. For estimating the length of a rock bolt one empirical formula state:

L= 1.4+ 0.184B

Where, L = bolt length B = span

For systematic bolting by tensioned rock bolts, a general rule is that the spacing between individual rock bolts should not be longer than the half of the bolt length. The support system as proposed in the Feasibility Report of the project is shown as Drawing No. 05. The machine hall and transformer hall caverns are proposed to be supported at crown and walls with 6m long 25 mm diameter rock bolts are staggered 1.75m centres and 100 mm thick shotcrete with welded wiremesh. The bus tunnels and the draft tubes are proposed to be

supported with 3m long 25 mm diameter rock bolts are staggered 2m centres and 100 mm thick shotcrete with welded wiremesh. The length of the rock bolt proposed, in the crown of powerhouse ( 6 m) seems to be on the optimum length of rockbolts. The bolt length from Q method and also from empirical methods stated in Table 10.7 stated above indicates the bolt length for crown and walls should be around 6m but spacing of rockbolts may be kept 2 m centres and welded wiremesh may not be required with shotcrete. In this study, the effect of rock support has not been included in the model. However, from the deformation patterns and principal stress plots, it is very clear that deformations in walls of the powerhouse cavern are much higher than the deformations at crown. Hence walls of the powerhouse are expected to cause more stability concerns than the crown. Thus in author’s view, the spacing of rock bolts provided at the crown and walls of machine hall and transformer chamber may be increased to 2m centre to centre. From the study of principal stress plots, it appears de-stressing is likely to occur on both the walls of the machine hall cavern especially between El 374 m to El 349m (below the crown and above the draft tube top junction). Hence longer rock bolts are necessary for walls of powerhouse rather than the crown. However the rock bolts provided on walls of the powerhouse are 6 m long. In any case, the length of rock bolt at walls should be more than that in crown. The use of longer rock bolts and higher pre-stressing of rock-bolts is recommended for the walls of the powerhouse as the deformations likely to occur at the walls are much more than at the crown, since the height of the powerhouse is much more than its span. Based on the empirical and classification methods and after simulation using the numerical model, it is found that the support system as proposed for the transformer chamber seems adequate. It is recommended that while designing the support system for large and complex underground works with multiple openings, no single method can be followed. The empirical and classification methods must be substantiated with numerical modelling. The likely failure mechanisms, development of tensile zones or de-stressing, excessive compressive stresses and problems related to weakness zones, fault zones and ingress of water must be studied with the help of geological investigations, mapping and numerical modelling before recommending the support system. No system should be followed blindly and a judicious mixing of the available system and information by all methods may be the best way.

CHAPTER 10

CONCLUSIONS & RECOMMENDATIONS

__________________________________________________________________ The Lower Himalayan region has mainly Phyllitic Quartzite rocks which may have good quality rocks along with intermittent poor rocks. The geological investigation of the tunnel area reveals that about 56 % tunnel length is in fair to good rock. The rest of tunnel is poor to fair rock. The Feasibility report of 510 MW Teesta Stage-V Hydroelectric project located in Sikkim, India has been prepared by National Hydro-electric Power corporation Ltd. The Norwegian experience could be applied to evaluate the design of the project. Different layout designs commonly used in Norway have been considered for design of tunnel, pressure shaft and power house cavern. The Location, orientation, shape and dimension of cavern have been optimised using Norwegian experience and using Empirical methods. Q-method has been used to evaluate and design tunnel, pressure shaft and power house caverns. The empirical design should be checked using a numerical model study to simulate the proposed design. Numerical modelling represents the most versatile computational method in the field of geo-technical engineering. The use of numerical methods is growing due to rapid increase in the field of computers. Geological problems are very complex, and hence very accurate modelling is very difficult and often not possible. Therefore the numerical models have to be simplified by making suitable assumptions and taking the most important aspects while neglecting the minor details. Numerical modelling is a very valuable tool in understanding the rockmass behaviour after excavation, the stress redistribution, studying possible failure mechanisms and failure zones and predicting approximate values of deformation likely. However, the results obtained by numerical modelling depends on the quality of input data. If the input data is reasonably good and representative then good quality results can be expected. If the quality of input data is not good then the results must be viewed judiciously. It is always recommended to carry out a back analysis whenever numerical analysis is being carried out. Instrumentation must be carefully planned and the results predicted by the numerical analysis should be compared with actual results obtained by the instrumentation. The actual deformations, as measured by instruments, should be used to refine the model for future use. It also gives confidence while carrying out numerical analysis in future. As discussed earlier, it is not possible to incorporate all minor geological features in the model. Hence while designing the support system the local geological features which have been ignored in the numerical model, must be accounted for and suitable measures be adopted to take care of them. The results of numerical modeling gives the basic guidelines of the likely behaviour of the rock mass and the results must be carefully used.

In this study the following conclusions have been made by carrying out the study based on Norwegian experience and empirical methods and further using numerical model analysis for different models:

1. The tunnel system could be realigned to reduce the total length of tunnel. The number of adits can be reduced from 5 adits as proposed in Feasibilty report to 3 adits, which can reduce the length of tunnel from 17.8 Km to 15.4 Km.

2. The tunnel could be designed as unlined tunnel with supports only at places, where they

are required. Geological investigations have revealed that about 56% of tunnel length pass in fair to good rock, which may not require any permanent support. The rock classification of tunnel length in the Feasibility report seems to be on the conservative side.

3. The tunnel excavation can be carried out using a Tunnel Boring Machine (TBM) after

carrying out laboratory tests of the rock samples in the proposed tunnel for selection of a TBM, suitable for tunnelling in the rock. It may reduce the construction period of project by about 2 years, which may improve the economic feasbility of the project.

4. Two or three numbers of small diameter tunnels can be bored using one TBM with a

time lapse, so that 1 or 2 units may be installed earlier and thus profit earned from sale of power produced from earlier installed units could be further utilised for the cost involved in installation of additional works.

5. The pressure shaft could be re-designed as unlined pressure shaft with concrete lining

limited in certain locations with shear zones and weak rocks. 6. The shape of the power house cavern should be decided based on stress situation and

quality of rock. Moderate stress field and good quality rock at power house location makes it possible to design a cavern with vertical walls and flat arched roof. The limited data available about the rock and stress situation makes it difficult to decide based on assumptions about stress situation.

7. Span of the cavern is the most critical dimension in power house design. Norwegian

experience for compact cavern layout designs could be used to optimise the cavern volume and reduce the cavern span. The span can be finalised based on turbine, generator and transformer manufacturer’s specifications for the project.

8. The power house cavern design based on empirical methods and Norwegian experience

should be verified using a numerical model study. 9. The Numerical model results indicate that the deformation of wall is much higher than

the deformation at the crown of the Machine hall as the height of the machine hall is more than two times the span.

10. The deformations in the walls of the transformer chamber is less as compared to machine hall as the ratio of height and span of transformer chamber is less as compared to the machine hall.

11. In transformer chamber the increase in wall deformation is less than crown deformation

due to excavation of bus tunnels and draft tubes. This is due to the fact that draft tube and bus tunnel excavation induces vertical deformation, which occurs at the crown. However in the machine hall, the deformation in both crown as well as wall increases almost in same ratio (on excavation of bus tunnels & draft tubes, i.e., model 2 & 4). This can be explained by the fact that many elements near the walls have already exceeded their permissible shear and tensile stress in the past and hence the excavation of bus tunnels and draft tubes leads to increase in deformation in walls of the powerhouse.

12. There is de-stressing on the downstream walls of the powerhouse cavern between EL 374.5m and EL 349 m. The major principal stress is tensile and it implies the loosened state of rock mass which must be strengthened and reinforced.

13. The de-stressed zone is slightly more in the in the model using Mohr-Coulomb’s

plasticity model than in the one using elastic material. The magnitude of the tensile stresses is also higher in the plastic model. This is due to the effect of non-linearity.Once the stress in the rockmass reaches the yield value it non longer takes any load and any further stress is distributed to the adjoining rockmass. It is author’s view that the plastic model is closer to the natural phenomenon if the values are chosen correctly.

14. From the deformation pattern and stress situation it is clear that walls of the powerhouse needs longer rock bolts than the crown. Hence it is recommended to use longer rockbolts in the walls and shorter in the crown.

15. The 6m long rock bolts at 1.75m centre on the crown and 100mmm thick shotcrete with

welded wiremesh of the caverns can be optimised based on rock excavated. If the cable anchors have to be used then they must be provided on the downstream wall of the powerhouse from El 374.5m to El 349 m. It can take care of loosening of rock mass and also prevent excessive deformations.

16. It is recommended to carry out a back analysis by comparing the actual deformations

with those predicted and refine the model.

17. It is also recommended to study the effect of support system by including it in the model.

REFERENCES

1 Barton, N. et al (1974) “ Engineering Classification of Rock Masses for the design of Tunnel Support ’’, in Rock Mechanics, Vol. 6, No. 4. 2 Bieniawski (1974). “ Engineering Classification of Rock Masses and Its

Application in Tunnelling’’, in Proc. 3rd ISRM – Congress, Denver 1974, Vol. IIa.

3 Broch, E. et al, 1995 ‘‘Design approach for Underground power plants’’ Hard rock Engineering, FHS, Oslo, pp 131-147.

4 Broch, E. et al, 1996 ‘‘Support of Large Rock Caverns in Norway’’ in tunnelling and

Underground Space Technology, Vol. II pp 11-19. 5 Desai, C. S. (1972). “Introduction to the Finite Element Method ’’, New York. 6 FLAC 3D User’s Manual Version 2.

7 Gunapala, L. (1997). “Evaluation and Redesign of Tunnels and Power house cavern at

Kukule Ganga Hydro power Project based on Norwegian experience’’ M.Sc. thesis in Hydropower Development, Department of Geology and Mineral Resources, Norwegian University of Science & Technology (NTNU), Trondheim.

8 Madan, M.M., (1996) “ Penetrating the Himalayas’’, in Tunnels & Tunnelling, June 1996. 9 Nilsen, B. and Thidemann, A. (1993). “Rock Engineering’’ , Hydropower

Development , Vol 9, Division of Hydraulic Engineering, Norwegian Institute of Technology, Trondheim.

10 Project Report (1998) ‘‘Feasibility Report on Teesta Stage-V Hydro-electric Project,

Sikkim’’, National Hydro-electric Power Corporation Ltd. Faridabad, India. 11. Project Report 1-88 (1988) ) ‘‘Hard Rock Tunnel Boring’’, NTNU, Trondheim, Norway. 12. Project Report 2-88 (1988) ) ‘‘ Tunneling, Prognosis for Drill and Blast’’, NTNU,

Trondheim, Norway. 13. Project Report 3-88 (1988) ) ‘‘Tunneling, Cost for Drill and Blast’’, NTNU, Trondheim,

Norway. 14. Sinha, R. S. (1989) “ Underground Structures Design and Instrumentation ’’ Denver. 15. Starfield, A. M. and Cundall, P.A., (1988) “ Towards a methodology for Rock Mechanics

Modeling’’, in Int. Journal of Rock Mechanics Mineral Science & Geomech. Abstr.

16. Saran, V.,(1999) “Three Dimensional Numerical Analysis of Underground works at Xiaolangdi Multipurpose Project in China’’ M.Sc. thesis in Hydropower Development, Department of Geology and Mineral Resources, Norwegian University of Science & Technology (NTNU), Trondheim.

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