Underwater Tunel Piercing

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CONTENTS PAGE

PREFACE 1

0 INTRODUCTION 3

1

METHODS OF PIERCING 1.0 Introduction 1.1 Open System 1.2 Closed System 1.3 Choice of Method

5 5 6 8 10

2

DEVELOPMENT 2.1 Existing Lake Taps 2.2 Outlook

11 11 12

3

SITE INVESTIGATIONS 3.0 Introduction 3.1 Investigations Methods 3.2 Exploratory Drilling 3.3 Visual Inspection 3.4 Choice of Piercing Site

13 13 14 16 17 18

4

TUNNELLING TOWARDS THE PLUG 4.0 Introduction 4.1 Probe Drilling 4.2 Georadar 4.3 Grouting 4.4 Smooth Blasting 4.5 Water Leakages 4.6 The Piercing Area 4.7 The Plug

19 19 20 23 26 29 31 32 35

5

THE BREAKTHROUGH ROUND 5.0 Introduction 5.1 Drilling 5.2 Charging 5.3 Coupling 5.4 Blasting

38 38 42 43 46 47

Continues next page

6

HYDRODYNAMICS 6.0 Introduction 6.1 Open System 6.2 Closed System 6.3 Explosives Gas Pressure 6.4 Retardation Pressure 6.5 Upsurge in the Gate 6.6 Transport of Mass 6.7 Model Testing

48 48 49 50 51 54 56 57 58

REFERENCES 59

APPENDIX A Previous Editions of the Report B Research Partners C Definitions and Parameters D Basis of the Report E Some Recent, Large Underwater Piercing in Norway F1 Seismic Methods F2 Consumption of Grouting Media in Systematically Fractured Rock Mass F3 Blasting with Short Rounds F4 Example of Application for Charging Plan F5 Example of Application for Upsurge in the Gate Shaft F6 Example of application for Explosives Gas Pressure G1 Example: Underwater Tunnel Piercing at Lake Fossevatn, Kobbelv Hydro Power Plant G2 The Troll Field Shore Approach

64 65 66 67 68

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76 80

87

95

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101 111

PREFACE

1

UNDERWATER TUNNEL PIERCING Project Report 19-99

The Project Report Series from the Department of Building and Construction Engineering at the Norwegian University of Science and Technology contains several reports about rock blasting, tunnelling and related subjects. The reports in the Project Report Series present updated and systematised knowledge on rock excavation and tunnelling to be used for: • = Economic dimensioning • = Choice of alternative • = Time planning • = Cost estimates, tender, budgeting and cost control • = Choice of excavation method and equipment. A list of available Project Reports may be requested from the Department of Building and Construction Engineering at NTNU The report is prepared by Amund Bruland, Gry Helle Nakstad, Steinar Roald and Professor Odd Johannessen. The project has been granted financial support by our external research partners, see list in Appendix B.

PREFACE

2

It is recommended that the references to this book should be made in the following way:

NTNU-Anleggsdrift (1999): Project Report 19-99 UNDERWATER TUNNEL PIERCING.

When copying from the report, the source should be stated. Trondheim, April 2000 Odd Johannessen, Professor Contact address: Amund Bruland Department of Building and Construction Engineering, NTNU N-7491 Trondheim NORWAY Telephone: (+47) 73 59 46 40 Fax: (+47) 73 59 70 21 E-mail: [email protected] Internet: http://www.bygg.ntnu.no/batek

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0. INTRODUCTION Background and Objectives

3

BACKGROUND AND OBJECTIVES A total of approximately 600 underwater piercings have been carried out in Norway during the last hundred years. In most cases these have been part of hydroelectric projects. The lake tap enlarges the water reservoir by establishing an intake level below the natural water level of the lake. Such reservoirs are lowered in the winter and refilled during the summer months.

Figure 0.1 Typical design of a Norwegian hydropower plant. Underwater tunnel piercing is also used in other areas such as preventing flooding, shore approaches for oil and gas pipelines from the North Sea, and may in the future be used for fjord or strait crossings by underwater tube bridges. The report gives a general introduction to the subject of submerged tunnel piercing. However, each break-through is unique and competent personnel with the necessary knowledge and experience in geology, blasting technique and hydraulics must be involved when planning and execute an underwater tunnel piercing.

0. INTRODUCTION Background and Objectives

4

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Figure 0.2 Underwater tunnel piercing - general layout. Variable factors may be: • = Water depth to the plug and the gate • = Tunnel cross section area • = Cross section area and volume of the plug • = Amount of explosives • = Distance between the plug and the gate • = Amount of sediments above and around the plug .

1. METHODS OF PIERCING 1.0 Introduction

5

1.0 INTRODUCTION Underwater tunnel piercings can be divided in two main categories: • = Closed system • = Open system. Open system piercing means that there is direct communication between the tunnel face, where the tunnel plug is, and the gate shaft. Closed system piercing means that the gate is located on the upstream side of the gate shaft. Hence, there is no communication between the plug and the gate shaft. This report deals mainly with the open type of piercing.

1. METHODS OF PIERCING 1.1 Open System

6

1.1 OPEN SYSTEM The principle of the open system piercing is shown in Figure 1.1. There is direct communication between the face of the tunnel, at the tunnel plug, and the atmospheric pressure in the shaft. The gate or the concrete plug is placed on the downstream side of the gate shaft. The tunnel between the rock plug and the gate is filled with water to avoid transportation of sediments towards the gate. When filling the tunnel system with water, the air cushion between the plug face and the tunnel will be compressed. To prevent air evacuation through the plug, the air pressure must be less than the hydrostatic pressure against the tunnel plug, see Figure 1.1.

H1 = water depth at the tunnel plug

H2 = air pressure at the plug, i.e. difference in level between the water in the gate shaft and at the plug.

If the water level of the reservoir is higher than the water level in the gate shaft, there will be a surge of the water in the gate shaft above the water level in the reservoir. This surge can be estimated. Normally, the water will rise 70 - 90 % of H3 above the level of the reservoir, see Figure 1.1. To avoid damage to the gatehouse, the following must be true:

H3 = difference in level between the water in the reservoir and in the gate

shaft H4 = difference in level between the gate house and the water in the reservoir c = constant (0.7 – 0.9).

Criteria 1: 12 HH < (m) [1.1]

Criteria 2: 34 HcH ⋅> (m) [1.2]

1. METHODS OF PIERCING 1.1 Open System

7

��

��

��

Figure 1.1 Open system piercing. Criteria 1: H2 > H1 to prevent the air being evacuated through the plug. Criteria 2: H4 > c⋅H3 to prevent surge of water into the gatehouse.

1. METHODS OF PIERCING 1.2 Closed System

8

1.2 CLOSED SYSTEM The closed type of piercing can be divided into two categories, depending on the water level in the tunnel system. Figures 1.2 and 1.3 show the principles of piercing with a dry tunnel and with a partly waterfilled tunnel. The revision gate is closed for both alternatives. If no precautions are taken, the high velocity water inflow can transport the rock debris from the plug far into the tunnel when the plug is blasted. When piercing from a partly waterfilled tunnel, the water inflow velocity and the transport of rock debris along the tunnel will be reduced. On the other hand, the maximum pressure in front of the gate will be increased. Providing a long tunnel between the plug and the gate can reduce the pressure.

��

Figure 1.2 Closed system piercing - dry tunnel.

1. METHODS OF PIERCING 1.2 Closed System

9

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Figure 1.3 Closed system piercing - partly waterfilled tunnel.

1. METHODS OF PIERCING 1.3 Choice of Method

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1.3 CHOICE OF METHOD Open system piercing gives low velocity water inflow and makes the trapping of the blasted rock possible. The estimation models for the highest surge of water in the gate shaft and the maximum hydraulic pressure against the gate during the plug blasting are satisfactory. On the other hand, the method is complicated regarding the filling of water and air into the tunnel. Thus, it may take long time from the coupling of the round until the blasting of the plug. The closed system is in most cases less complicated and less expensive to perform, especially with a dry tunnel. There is some uncertainty concerning the transportation of rock debris from the plug, which is the main disadvantage of the method. However, the rock debris is normally fine grained due to the overcharging of the final round, and the problems are estimated to be small. A closed system piercing is normally not suited for very short tunnels.

OPEN SYSTEM CLOSED SYSTEM

Advantages • = Simple hydrodynamic conditions.

• = Controllable trapping of the blasted rock.

• = Maximum pressure against the gate may be estimated with low uncertainty.

Advantages • = Easy to perform.

• = Short time from coupling of the round to blasting of the plug.

Disadvantages • = Complicated to perform.

• = Long time from coupling of the round to blasting of the plug.

• = Comparatively high costs.

Disadvantages • = Complicated hydrodynamic conditions.

• = The rock debris may damage the gate.

• = Requires a long tunnel between the plug and the gate.

Table 1.1 Comparison between open and closed system, summarised from

experience.

2. DEVELOPMENT 2.1 Existing Lake Taps

11

2.1 EXISTING LAKE TAPS There are a large number of lakes in the mountain regions of Norway. Hundreds of these are utilised as reservoirs for hydroelectric power plants, by tapping the water below the normal water level, see Figure 0.1. The first Norwegian lake tap we know of, was done in 1893-94. Around the turn of the century, lake taps were performed to provide reservoirs for hydroelectric power plants. The exact number of lake taps in Norway is not known. Most probably 600 lake taps have been performed. The variation in water depth is from a few metres to more than one hundred metres. The thickness of sediments above the plug has been between 0 m and 6 m and the plug length has varied between 1 m and 10 m. The cross section of the plug has varied from 1 m2 to 95 m2 and the tunnel length beneath the lake bottom varies from a few metres to approximately 1000 m. The water leakage through the plug has been from less than 1 l/s to several hundred litres per second. Some examples of deep lake taps are given below: • = 105 m Lake Juklavatn, the Folgefonn Hydropower Project, 1974. • = 109 m The Blåsjø Reservoir, the Ulla-Førre Hydropower Project, 1986. • = 120 m Lake Fossvatn, the Kobbelv Hydropower Project, 1987. • = 170 m Troll Shore Approach, where gas pipelines from the North Sea was

pulled in through the piercing, 1994. Experience from the early period of lake taps together with measurements and the analysis of later lake taps have provided the basis for the know-how of Norwegian consultants and contractors in this field.

2. DEVELOPMENT 2.2 Outlook

12

2.2 OUTLOOK Most lake taps are made to increase the reservoirs for hydroelectric power plants. Today, underwater tunnel piercing is also a possible solution for the shore approaches of oil or gas pipelines, power transmission cables and also for strait crossings by underwater tube bridges.

Figure 2.1 Tunnel piercing to connect rock tunnel and a submerged tube bridge.

3. SITE INVESTIGATIONS 3.0 Introduction

13

3.0 INTRODUCTION Underwater investigations are generally very expensive. Due to this, it is reasonable to divide the investigation program into steps to limit the total costs of the investigations. Unfavourable piercing sites may then be rejected at an early stage of the planning, without using resources in further investigations. Sediments Depositions of sediments at the lake bottom cause problems for the piercing. The nature and thickness of the sediments must be investigated. Where there is minimum risk of landslides, a thin sediment cover may be advantageous. If the rock in the piercing plug area is fractured, the sediments will help filling the open joints and reduce the leakage of water into the tunnel. It may be very difficult to handle scree material containing large blocks or compacted sediments. Geological Mapping The planning of underwater piercings requires precise investigations in several phases to collect relevant topographic and geological information such as the distribution between rock and sediments and a general view of the joint systems and fractured zones in the area.

3. SITE INVESTIGATIONS 3.1 Investigation Methods

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3.1 INVESTIGATION METHODS Seismics There are two main types of seismic investigation methods: • = Reflection Seismics • = Refraction Seismics. Seismic investigations must be carried out and interpreted by specialist. For underwater piercings, the main purpose of the seismic investigations is to find the depth to solid rock, i.e. the thickness of sediments at the lake bottom. It is also of interest to obtain information about the sediment types, e.g. clay, moraine, scree, etc. In some cases it is also possible to find fracture or fault zones by using seismic methods. Hence, the piercing may be located outside such zones. See also Appendix 3.1 Reflection Seismics. Georadar The georadar is an electromagnetic instrument. The instrument is relatively new and has so far not been used extensively in Norway. Testing and use in other countries have shown promising results. Georadar may be used in the preliminary investigations for underwater piercings to find the depth of water and the thickness and extent of the sediments on the bottom. Georadar may only be used in freshwater, and it does not penetrate deeper than 25 - 35 m below the water surface. For tunnel excavation towards the plug, it is possible to use Georadar in mapping and to a certain extent for the description of fractured zones, joints with water, etc. from within drillholes. The drillhole Georadar system will be discussed in Section 4.2. Using a Georadar to find the depth to rock is limited by the range of the Georadar (25 - 35 m in freshwater). The range of the Georadar is determined by the conductivity of each material. Long range is obtained in insulating materials, e.g. gravel and sand. Good conductors, e.g. clay and seawater, severely limit the range.

3. SITE INVESTIGATIONS 3.1 Investigation Methods

15

Material Maximum Range (m)

Rock 75 - 100 Gravel and sand 60

Freshwater 20 - 35 Peat 15 - 20

Moraine 15 Silt 5 - 10

Clay 2 - 4 Table 3.1 The range of the Georadar in varying materials. The georadar cannot be used to find the thickness of sediments at great depths or in seawater. Costs One day of measurements with a georadar costs approximately NOK 30 000.

3. SITE INVESTIGATIONS 3.2 Exploratory Drilling

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3.2 EXPLORATORY DRILLING

3.21 Percussive Drilling Percussive drilling is a method for finding the depth to solid rock and the quality of the rock at the piercing site. The method is based on registering the drilling rate and the torque as the drill bit penetrates sediments and solid rock. The method is mainly used in cases where refraction seismics do not give a clear interpretation. In general a probe hole should be drilled at least 2 m into solid rock to ensure penetration of large blocks. At normal depths of water, percussive drilling is carried out from a barge or a raft. Costs At normal depths, the costs are approximately NOK 1000 – 2000 per m of drillhole.

3.22 Core Drilling The preliminary investigation methods described above only give an indication of the types of material at the site. For refraction seismics, fractured rock gives lower seismic velocities than homogenous rock, and for percussive drilling the penetration rate is higher for fractured than homogenous rock. When accurate information is required about the rock quality in fractured zones and weak rock types regarding e.g. leakage, swelling clay and degree of fracturing, core drilling is a solution. With core drilling it is possible to drill holes of several hundred metres. The core drilling may also be equipped for directional drilling to increase the accuracy of long holes. Core drilling is expensive and should only be used when the required information cannot be found by other methods. Costs When core drilling from an onshore site, the costs are approximately NOK 1000 per drilled metre. Directional core drilling has a cost of approximately NOK 2000 per drilled meter.

3. SITE INVESTIGATIONS 3.3 Visual Inspection

17

3.3 VISUAL INSPECTION A diver or a submarine vehicle (ROV) may do a visual inspection. Documentation of the observations is usually photographs or videotapes. The test sampling of sediments is optional. At water depths down to 50 m, it is normal to use divers. If the conditions are favourable, diving is an inexpensive and satisfactory investigation method for the piercing site. At greater depths, the efficiency of divers is lowered due to the long time required for decompression. Investigations by ROV (Remote Operated Vehicle) give a very good impression of the conditions at the piercing site. Close-up photographs and videotapes can be very detailed, even in 3D. However, it may be difficult to get a general view of the formations. The ROV is also well suited to investigate the break-through area and the tunnel system after the blasting.

3. SITE INVESTIGATIONS 3.4 Choice of Piercing Site

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3.4 CHOICE OF PIERCING SITE The general demands for a suitable piercing site are always the same. The piercing site must be located where the rock quality is favourable and the thickness of sediments is limited. Some sediment cover (0.5 - 1.0 times the diameter of the piercing plug) will usually improve the grouting conditions.

4. TUNNELLING TOWARDS THE PLUG 4.0 Introduction

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4.0 INTRODUCTION The excavation of the final part of the tunnel towards the piercing plug is a very demanding task for the tunnelling crew and the site management. In addition to favourable rock mass quality, an experienced tunnelling crew is of vital importance for the time required and the excavation costs. Only one tunnelling crew should take part in the excavation of the last part of the tunnel. The time consumption and costs for excavation of the final tunnel stretch towards the plug may be estimated using the project reports published by the Department of Building and Construction Engineering at NTNU, Trondheim. • = Project Report 2A-95 TUNNELLING Blast Design • = Project Report 2B-95 TUNNELLING Prognosis for Drill and Blast • = Project Report 2C-95 TUNNELLING Costs for Drill and Blast

4. TUNNELLING TOWARDS THE PLUG 4.1 Probe Drilling

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4.1 PROBE DRILLING

4.11 Standard Procedures On the basis of the preliminary investigations, the tunnelling operations continue according to standard procedures until there is 50 - 100 m of the tunnel remaining until the piercing plug area. The purpose of probe drilling is mainly to secure the next round, regarding the location of fracture zones and water leakages, and to make sure that the thickness of rock between the tunnel and the reservoir is sufficient. The length of the probe holes should be adapted in such a way that there is an overlap of at least one round length between consecutive probe drilling cycles. If there is poor quality rock at the face of the tunnel, the probe holes should be drilled with care to avoid drainage of too much water into the tunnel. The positioning and direction of the probe holes may vary. To get the best results from the probe drilling, it is necessary to develop routines for leakage measurements, registration of clay in the flushing water, and variations in the drilling rate due to joints and fracture zones. It is important to avoid blasting the next round before the probe holes are evaluated and possible precautions are made, e.g. grouting, reduced round length or change of tunnel direction. When the distance from the tunnel face to the bottom of the lake is 12 - 15 m, several holes should be drilled into the lake to map the surface and confirm and supplement the seismic profiles. The holes should reach at least 1 - 1.5 m into the water or the sediments to avoid misinterpretations of e.g. a fracture zone or an open joint. After carefully tunnelling towards the piercing plug and before stopping the tunnel roof, probe drilling is performed in and around the plug to map it and to place the cut in the best rock conditions. Finally, an adjusted drilling pattern for the plug is designed.


21

Holes that are drilled into the water must be sealed from inside the tunnel. Wooden plugs may be used for water pressure up to 40 - 50 m. Drillhole packers must be used for higher pressure. The quality of the rock mass and the water leakage determine whether a hole has to be grouted or not. However, the hole must be sealed at the face of the tunnel to avoid that the hole is charged by a mistake later.

4.12 Example of Application It must be emphasised that the probe drilling should have an adapted design for each piercing. The following example is only presented as an illustration. The extent of probe drilling depends on the rock mass quality. Besides, supplementary holes must be drilled in case of hitting fracture zones, marked single joints, etc. The purpose is to find the extent and orientation of the zones or joints to be able to seal and grout the tunnel and the plug in the most favourable way. If necessary, the probe holes may be used as grouting holes. The holes must satisfy the length and direction requirements given in Section 4.3. The final stretch of approximately 50 m of the tunnel towards the plug is divided into three phases. In Phase I the probe drilling is done every third round. The overlap is approximately the length of one round. In Phase II the round length is reduced, and drilling into the water is done to ensure safety at the tunnel face. The general distance to the reservoir is most often interpreted from the seismic investigations. The upward and forward distance to the reservoir must be verified by the probe drilling.


22

In Phase III the probe holes are relatively short to avoid large water leakages. In this phase it is important to blast very carefully to prevent the opening of joints towards the reservoir. If necessary, the round length may be reduced more than in Phase II. When the tunnel is finished, drilling through the plug and into the water is done to map the plug accurately. Finally, the piercing area of the tunnel is stopped in such a way that the shape is favourable, see Sections 4.6 and 4.7.

��

�� Figure 4.1 Suggestions of how to do the probe drilling for the final stretch towards

the piercing plug.

4. TUNNELLING TOWARDS THE PLUG 4.2 Georadar

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4.2 GEORADAR As mentioned in Section 3.2, georadar may also be used during the excavation of the tunnel towards the plug. The georadar is placed in a drillhole drilled from the face of the tunnel. The purpose of these drillholes is to locate and characterise fracture zones, marked single joints etc. in the rock mass in front of the tunnel. The preliminary investigations have enabled mapping of the piercing site from the "outside" by e.g. refraction seismics. The same area is also mapped from the "inside" using georadar. This verification of the preliminary investigations increases the quality of the plug mapping considerably. The principle of the georadar used in a drillhole is the same as described in Section 3.2. The positioning of the receiver is optional. It may be placed in the same hole as the transmitter or in another hole. There are two main types of reflectors • = Planar reflectors, e.g. fracture zones. • = Point reflectors, e.g. a drill hole. When using the georadar system, the distance from a reflecting object is determined by measuring the difference in arrival time between the direct and the reflected pulses.


24

Figure 4.2 Principle for the use of georadar in a drillhole. The results appear in a diagram where the positions of the receiver and the transmitter are shown along one axis and the distance between the drillhole and the reflector (travel time) along the other axis. As the georadar is moved along the drillhole, the distance to the reflector will vary and show a characteristic pattern in the diagram • = Planar reflectors show two lines formed as a V. • = Point reflectors show a hyperbola. It is also possible to measure the strength of the reflection and to some extent this may describe the fracture zone or the joint. For instance, a joint filled with clay or another good conductor usually gives a strong reflection, see Table 3.2 in Section 3.4.


25

The range of the georadar is dependent on the rock type and the quality of the rock mass. Under favourable conditions (e.g. some granites and gneisses) fracture zones may be detected as far as 100 m from the drillhole. Georadar may also be used in cross-hole measurements, which means that the receiver and the transmitter are placed in different drillholes. The velocity and the amplitude of the direct pulse between the transmitter and the receiver give information about the rock mass quality between the drillholes. This method is most suitable in vertical drillholes drilled from the surface. Cross-hole measurements may cover large areas. If the quality and the type of rock are favourable, the distance between the drillholes may be up to 250 m. Restrictions Dipole antennas are used because of the drillhole diameter. This means that the pattern in Figure 4.2 has cylindrical symmetry. Consequently, it is not possible to obtain the orientation (dip and strike) of a fracture zone by measurements in a single drillhole. Hence, while the angle of intersection between a drillhole and a zone may be decided, the zone may still rotate along a conical surface. When deciding the dip and strike of a zone, the measurements from at least three different holes have to be combined. Development of directional antennas will make it possible to decide the orientation of the zone from measurements in a single drillhole. There is a limitation concerning the drillhole diameter. The diameter of the georadar used in Sweden was 56 mm, while the standard drillhole diameter in tunnelling is 45 mm or 51 mm.

4. TUNNELLING TOWARDS THE PLUG 4.3 Grouting

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4.3 GROUTING

4.30 Introduction Grouting is the only adequate method to stop leakages in the rock mass. Normally, grouting also stabilises the rock mass. The choice of grouting medium is dependent on water quantity, water pressure, the joint structure, maximum leakage allowed, curing time, etc. There are two main types of grouting media • = Cement-based suspensions • = Chemical media.

4.31 Cement-based Suspensions The cement-based suspensions are in most cases the best alternative. During the grouting, joints and fissures are filled with solid particles. Because of this, cement grouting is hardly relevant when the joint aperture is less than 1 - 2 mm. The standard procedure is to start with a relatively thin suspension, e.g. a mix of water and cement in a ratio of 3 to 1 (by weight), and gradually make the suspension thicker as the pressure increases. Cement-based grout with very low setting time is now commercially available.

4.32 Chemical Grouting Media There are many types of chemical grouting fluids or solutions. These are free of particles. Hence, the chemical media have the best infiltration capacity in thin fissures. The hardening time is usually shorter than for cement-based suspensions. Chemical grouting is seldom used for tunnel piercing.


27

4.33 Pregrouting All grouting should be performed as pregrouting i.e. grouting ahead of the tunnel face before the tunnel reaches the leakage area. The acceptable amount of water leakage into a water tunnel is dependent on several factors. In rock mass with no clay-infected joints or fractures, it is the conditions at the face of the tunnel during the drilling and charging of the round and the water pumping capacity which limit the acceptable water leakage. Caution has to be taken when the rock mass is clay-infected. Even the smallest leakage of water containing clay, indicates that grouting is required. If the clay in the joints is washed out, the leakage may increase considerably. If there is a larger clay-infected zone, it may be necessary to drill drainage holes outside the grouted area to reduce the water pressure until the rock mass is stabilised. The grouting holes are drilled systematically, but their length, position and number may vary. The holes are mainly placed in the contour and are arranged like an outward fan. The centre of the face should also be grouted. The grouting pressure must be higher than the water pressure, but not so high that the rock mass will be split or lifted. If it is difficult to achieve the required tightness due to too low grouting pressure, drilling of more grouting holes is preferred rather than increasing the pressure. The use of a complete grouting screen may vary. In order to ensure that the rock surface can withstand the grouting pressure, the water should not come closer than 2 m from the contour. This is very important when the rock mass is of poor quality at the face. The grouting screen is usually efficient 8 - 10 m in front of the face. To achieve good sealing, the grouting media must penetrate a minimum distance from the tunnel to avoid the grouting media disappearing when blasting relieves the joints. The range of the grouting screen may vary, but as a rule of thumb a distance corresponding to the width or diameter of the tunnel should be grouted as a minimum. Control holes must be drilled when the grouting is completed. The length of these holes should be equal to the length of the grouting holes.


28

Experience shows that being prepared for grouting is a necessary precaution. Equipment and grouting media should be immediately available. If the complete cross section of the tunnel reaches a leakage zone, it is more complicated to bring the leakage under control, and time and cost estimates may be endangered.

4.34 Consumption of Grouting Media It is very difficult to predict the amount of grouting media necessary, due to the fact that jointing or fissuring of the rock mass is most often not systematic and therefore difficult to map. It is important to have sufficient grouting media available. Otherwise, if the grouting has to stop due to lack of grouting media, before sufficient pressure is achieved, the grouting is wasted. When classifying the fracturing of the rock mass, distinctions are made between: 1. Systematically fractured rock mass

- parallel oriented joints and fissures - foliation planes or bedding planes

2. Marked single joints 3. Small fracture zones 4. Crushed zones. The classification system is taken from Project Report 1D-98 HARD ROCK TUNNEL BORING Geology and SiteInvestigations. An example of how to estimate the consumption of grouting media in systematically fractured rock mass is given in Appendix F2. An example of cost estimation for grouting operations is also included.

4. TUNNELLING TOWARDS THE PLUG 4.4 Smooth Blasting

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4.4 SMOOTH BLASTING

4.41 Performing the Blasting When the tunnel excavation approaches the piercing area, the strain on the remaining rock must be reduced as much as possible. This is done by: • = Reducing the confinement of the holes. • = Reducing the amount of explosives per detonator no. Methods to reduce the confinement are: • = Drilling the contour holes approximately 50 cm shorter than the other holes. • = The row next to the contour is drilled parallel to the contour, see Figure 4.8. • = The cut is drilled slightly longer than the rest of the round. Methods to reduce the amount of explosives per detonator number are: • = Reduce the round length gradually. If the quality of the rock is poor, round

length down to 1 m may be necessary. • = Blasting with a double row of contour holes dividing the round in two, i.e. blast

half the round and then stope the rest. • = Use all the available detonator numbers.

4. TUNNELLING TOWARDS THE PLUG 4.4 Smooth Blasting

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a) Side view.

b) Top view. Figure 4.3 Methods to reduce the confinement: - The contour holes are drilled 0.5 m shorter than the other holes. - The row next to the contour is drilled parallel to the contour.

4. TUNNELLING TOWARDS THE PLUG 4.5 Water Leakages

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4.5 WATER LEAKAGES The most efficient way to seal leakages is pregrouting. It is very difficult to stop leakages that already exist at or behind the face. Such an uncontrolled leakage may occur when a non-grouted joint is opened up because of the blasting. However, if an uncontrolled leakage occurs, there are some methods to be tried: • = Wedging • = Cover the face with shotcrete • = Drilling of relief holes • = Grouting with low setting time. Wedging means placing of e.g. wooden plugs where the leakages occur. The method is possible for low water pressure and if the leakages are located in a specific hole or in specific joints. Wedging is not suitable if the face is very fractured and is leaking "everywhere". Using shotcrete to seal the leakages is not recommended. It is difficult to obtain adhesion between the shotcrete and the wet rock surface. Relief holes may be a solution for high water pressures. Holes are drilled from the roof and the walls behind the face. This will reduce the water pressure and the water flow to facilitate wedging or grouting. Grouting with low setting time may be used in cases where the water flows from joints and crushed zones of some size.

4. TUNNELLING TOWARDS THE PLUG 4.6 The Piercing Area

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4.6 THE PIERCING AREA When using an open type of piercing, the level of the tunnel at the final stretch towards the plug must allow the plug itself to be at the end of a shaft that is at least 5 metres long. The shaft is needed for an air pocket under the plug when the tunnel is filled with water. The shaft should be directed perpendicularly to the bottom of the lake to make the drilling and charging of the breakthrough round easier. Before the drilling of the breakthrough round starts, a detailed probe drilling programme must be carried out to map the tunnel plug precisely and place the cut as favourably as possible. The plug is mapped as shown in Figures 4.4 and 4.5 by drilling into the water. The number of drillholes should be determined in each case. Blasting of the rock pit under the plug should be done before the stopping towards the plug starts. To reduce the strain on the plug itself as much as possible, only the mucking out of the pit should remain before the stopping starts. The placing and the size of the pit depend on the type of piercing that is to be performed, the total volume of the plug and the assumed thickness of sediments covering the plug. See also Section 6.6.


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Figure 4.4 Mapping of the tunnel plug and placing of the probe holes. Side view.


34

Figure 4.5 Tunnel plug and position of the probe holes.

4. TUNNELLING TOWARDS THE PLUG 4.7 The Plug

35

4.7 THE PLUG From a construction point of view, the plug should be approximately square with broken corners, as shown in Figure 4.5. Broken corners cause a reduction of the confinement, and the square shape of the plug makes systematic drilling possible with an even distance between the drillholes. It also makes the control of the drill holes easier. The opening of the lake tap should be smallest near the bottom of the lake. Possible large boulders or blocks will then be caught in the plug before jamming the shaft. A simple rule based on experience is often used when deciding the thickness of the plug:

tp = plug thickness wp = plug width, i.e. the shortest side of the plug cross section (m) Depending on the water pressure, the cross section of the plug and the quality of the rock, the thickness of the plug should range from 4 to 8 m. The width of the plug should not exceed 6 to 7 m. A general design of the piercing area and the tunnel plug is shown in Figure 4.6.

(m) 2.1 pp wt ⋅= [4.1]


36

Figure 4.6 The piercing area and the tunnel plug. Side view.


37

�

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Figure 4.7 The plug thickness as a function of the plug’s cross section.

5. THE BREAKTHROUGH ROUND 5.0 Introduction

38

5.0 INTRODUCTION Empirical data from some piercing done in Norway in recent years are shown in Figures 5.1, 5.2 and 5.3. • = Specific charging as a function of the plug cross section. • = Specific drilling as a function of the plug cross section. • = Necessary areas of large holes as a function of the plug cross-section. In the following example of application, the above mentioned graphs are used. Due to the relatively large part of large holes in the piercing plug, we have chosen to include the drilling of the large holes in the specific drilling, which is not the case for specific drilling in regular tunnel blasting. See Project Report 2A-95 TUNNELLING Blast Design.


39

�

�

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��

�

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Figure 5.1 Specific charging for the piercing plug.


40

��

��

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Figure 5.2 Specific drilling for the piercing plug, including the large holes.


41

��

��

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Figure 5.3 Necessary area of large holes.

5. THE BREAKTHROUGH ROUND 5.1 Drilling

42

5.1 DRILLING A satisfactory result depends on precision mapping of the plug and accurate drilling of the holes in the plug. The holes should be drilled with a guiding tube on the drill rod to achieve the required accuracy and prevent the holes from intersecting each other. This is most important for the cut. The tunnel crew that has excavated the final stretch of the tunnel should also drill the holes of the plug. This means that the tunnel crew should be familiar with the rock conditions, reducing the possibility of misunderstanding or misinterpretation of joints, leakages etc. The holes of the plug are drilled until 0.5 m of solid rock remains towards the lake. The surface of the rock and the thickness of the plug must be known through the probe drilling into the water and from the preliminary investigations. When each hole is drilled, the hole-depth is measured and noted. After drilling all the holes, before charging, the length of the holes must be checked. Holes that have water leakages or are filled with sediments have to be lined. These holes must be drilled to a larger diameter to make enough space to use a liner, such as plastic tubes. The inner end of the tubes must be sealed. Lining of all holes may be necessary when there is a long time between the drilling and the charging of the plug round. The drilling pattern must be as simple as possible with an even distance between the holes. This makes the drilling operations easier to control. The cut is placed in the most favourable area with regard to rock quality and confinement. Good rock quality gives precise drilling and prevents the holes to intersect. The confinement is lowest where the plug is at its thinnest. The cut is made by drilling the large holes and extra charged holes between the other evenly distributed holes. A more detailed description is given in Appendix F4. If the tunnel plug is very thick, the holes of the cut should not be placed too close, to avoid intersection. The most used diameters are 45, 51 or 64 mm for the charged holes, and 76 or 102 mm for uncharged large holes in the cut.

5. THE BREAKTHROUGH ROUND 5.2 Charging

43

5.2 CHARGING For all tunnels piercing, the explosives and the detonators must be water and pressure proof for the time between charging and blasting at the current depth of water. The following rule of thumb is used to find the amount of explosives to be used in the plug: • = Twice the normal consumption in a tunnel of the same construction area. • = Add 0.01 kg/sm3 in addition per metre of water depth. Finally, 10 % is added for unforeseen loss of holes. There are three methods used to charge a hole: • = Charging directly in the drill hole. • = Lining the drillholes with e.g. plastic tubes, and then charging the hole. • = Put the explosives into plastic tubes in advance, and then put the tubes into the

drill holes. The drillholes must have an uncharged length of about 0.5 m. When a drill hole is charged, a stemming plug is placed to keep the charge from falling out. Finally, a conical wooden plug with a slot is placed to seal the hole. The initiating cables are carried through the slot, see Figure 5.4. There must be two detonators of the same number in each hole. The detonators are placed in cartridge number three from top and bottom of the hole. For a thin piercing plug with short holes, the detonators may be placed in the second cartridge from top and bottom of the hole.


44

Figure 5.4 A charged drillhole. In regular tunnel blasting, the detonators are placed in the bottom of the hole. For a piercing plug, there are two throw directions. Hence, the best effect of the explosives is achieved by placing the detonators at some distance from the top and the bottom of the hole. For piercing plugs, electrical detonators Group 1 are normally used. The initiating cables must be protected from being damaged during the charging by a flexible tube. Non-electrical detonators may also be used, but are normally not recommended. The use of electronic detonators is very promising for tunnel piercing. The reason for using electrical detonators is that the function (i.e. resistance) of each detonator, the firing cable and finally the total round may be checked before blasting.


45

Group 1 detonators require low voltage to be initiated. This means that the protection against unintended initiation because of e.g. atmospheric electricity (lightning) is low. According to this, slow detonators may be used as a precaution measure. Slow detonators require considerably higher voltage to be initiated. However, using slow detonators also involves disadvantages such as: • = High voltage blasters are heavy and inconvenient to handle. • = High voltage requires good insulation of the cables and the connections. • = When slow detonators are used in water, the high voltage increases the

possibility of "current leakage" and unsuccessful blasting. Considering the disadvantages of slow detonators, the standard choice is detonators from Group 1. The detonators should be designed to resist the actual water pressure for the time required before firing of the round. The following precautions should be made during charging and of coupling the round to achieve satisfactory safety: • = A lightning detector must be placed in a visible position in the plug area. • = No electric cables in the plug area. • = Light is only provided by headlamps. • = Water pumps are driven by compressed air.

5. THE BREAKTHROUGH ROUND 5.3 Coupling

46

5.3 COUPLING In each drillhole there are two detonators of the same number. The detonators are coupled in two separate series, each of these are also coupled to a separate connecting cable. Then the two series are parallel coupled to a new, double connecting cable. To ease control, the inner and outer detonators in each hole are placed in their own series. This presupposes that the detonators are marked (e.g. red and green tape). The marking should be done before bringing the detonators to the piercing area. The initiating cables must not be cut because the resistance in the two parallel series has to be equal. It is required that the difference in resistance between the two series does not exceed 5 %. If the difference is larger, there is a risk that only one of the detonator series is initiated. During coupling, the series must be checked in intervals with a certified ohmmeter. The working conditions in the piercing area are often difficult, with a constant inflow of water. Hence, the coupling has to be simple and safe. Normally manual coupling with tape is recommended. Special coupling clips may also be used. As mentioned, the detonator cables must not be cut. Any extra cables are to be rolled up and fastened to plugs in the roof by strong tape, securing the cables from being damaged during the disassembly of the scaffolding. The connecting cable must also be fastened to the roof of the tunnel to avoid damage. The place where the firing is done is usually the gatehouse. The connecting cable is carried through the gate shaft from the tunnel. The cable may be carried through a tube to protect the wire in the shaft. In water filled or semi water filled tunnel, objects that will float must be removed from the tunnel to avoid possible damage to the connecting cable when the tunnel is filled with water. The coupling of the connecting cable must be done carefully with clips that correspond to the diameter of the cable. The connections must be well insulated and checked. Checking current leakage and insulation damage is performed with a special instrument that measures the ground resistance of the detonator circuit. Damaged detonators have to be replaced and damaged insulation has to be repaired.

5. THE BREAKTHROUGH ROUND 5.4 Blasting

47

5.4 BLASTING Before blasting, one last measurement by the ohmmeter is done to check that the coupling and wiring is all right. The blasting machine is chosen with regard to: • = Type of detonators and detonator group used • = Number of detonators • = The resistance of each detonator • = The resistance in the connecting cable • = Type of coupling (series/parallel).

6. HYDRODYNAMICS 6.0 Introduction

48

6.0 INTRODUCTION The underwater piercing is one of the most critical phases during the construction of a hydropower plant. The blasting of the piercing plug will generate loads on the different structures that exceed the static and dynamic loads experienced during operation. For structures close to the piercing area, the loads generated by the blasting will be the design loads. The hydrodynamics of a tunnel piercing is very complex, with non-stationary volume flow and compression and three-phase flow of water, blast fumes, air, broken rock and sediments. A closed tunnel piercing is more complex than an open piercing, as far as theoretical modelling, calculation and laboratory testing are concerned. From a construction point of view, a closed tunnel piercing is preferred. Experts must do the measuring of pressure during the blasting of the plug. This particularly relates to pressure against the gate, pressure in the air pocket under the plug and finally, the pressure in the gate shaft for open piercing. This chapter contains two examples of application. These are examples of how to use the calculation models developed at The River and Harbour Laboratory at SINTEF and NTNU, Trondheim. During the planning and execution of underwater piercing, experts on hydrodynamics, geology and rock blasting should be consulted.

6. HYDRODYNAMICS 6.1 Open System

49

6.1 OPEN SYSTEM The method requires certain water filling of the tunnel to reduce the transport of rock and sediments towards the gate and limit the surge of water in the shaft after the blasting. The level of water filling determines the design pressure. For a high level of water filling, the tunnel from the gate to the plug is filled with water except for a certain volume of compressed air, an air pocket, below the plug. See Figure 6.3. Low or no water filling requires a minimum length of tunnel between the gate shaft and the plug. High water filling should be used for short tunnels. Normally, the pressure in the air pocket should be at least 10 mWc (= 1 bar) less than the outside water pressure at the plug. The design pressure, the explosives gas pressure, is the increase of pressure which occurs when the explosives gas from the blasting is conveyed to the air pocket. Without this pocket of air, the pressure shock from the explosive gas will pass through the water towards the gate with little loss of amplitude and there will be a considerable possibility of causing damage. The water filling cannot be started until the piercing round is charged and coupled.

6. HYDRODYNAMICS 6.2 Closed System

50

6.2 CLOSED SYSTEM With a high level of water filling, the design pressure in a closed system is the same as for an open system. A low level of water filling means that the tunnel is less than half filled by water, 20% water filling is standard. In this case the design pressure is generated by the inflow of water through the blasted plug. The water compresses the air in the tunnel and the increasing air pressure (the retardation pressure) reduces the inflow. The maximum pressure occurs when the water inflow stops. The pressure in the compressed air in the tunnel is higher than the external water pressure and the water starts to flow out of the tunnel. Types of piercing and correspond pressure calculations (design pressure) are shown in Figure 6.1.

Figure 6.1 Type of piercing versus critical pressure.

Open Surge in shaft Closed with highwater level Explosives gas pressure Closed with low water level Retardation pressure

6. HYDRODYNAMICS 6.3 Explosives Gas Pressure

51

6.3 EXPLOSIVES GAS PRESSURE When the explosives in the piercing round are detonated, they will generate an explosives gas pressure which is dependent on the amount of explosives in the round and the pressure and volume of the air pocket. At detonation, 1 kg of explosives develops approximately 0.8 Nm3 of gas (1 Nm3 is the volume of gas which makes up 1 m3 at 1 bar and 0°C). The explosive gas is conveyed to the air pocket and causes a sudden increase in the pressure. The pressure increase will be transmitted through the water as a pressure wave towards the gate, where it will be reflected. The pressure increase at the gate is twice the amplitude of the pressure wave. The calculation presupposes the following: • = All the explosives gas is conveyed to the air pocket. This happens so fast that

the air pocket is said to be of constant volume during the blasting. • = There is no loss of energy in the pressure wave transmitted towards the gate. Variables included in the calculations are: p0 = the pressure in the air pocket before detonation pg = maximum pressure in the air pocket after detonation p1 = static pressure against the gate before detonation pd = design pressure against the gate. The pressure increase in the air pocket at firing is:

When the pressure wave hits the gate and is reflected, the increase of pressure is:

0 ppp ge −= (bar) [6.1]

)(2 0ppp gw −⋅= (bar) [6.2]


52

Design pressure against the gate: pd = p1 + 2(pg - p0)

To achieve safe results from the calculations, the input data must be correct (e.g. amount of explosives, pressure and volume of the air pocket): • = The amount of explosives may be checked during charging. • = The volume of air below the plug may be checked during the water filling. • = Methods to check the size of the air pocket include:

- Using a number of electrical circuits on a vertical column. As the water level rises, the circuits in known positions are closed. - A separate hose is installed from the plug area to the gate shaft, connected to a tank with a known level of overflow. The difference between the overflow level and the water level H may be read off in the gate shaft as the difference between the water level in the shaft and the hose.

Both methods are illustrated in Figure 6.2.

)(2 01 pppp gd −⋅+= (bar) [6.3]


53

Figure 6.2 Methods to check the size of the air pocket.

6. HYDRODYNAMICS 6.4 Retardation Pressure

54

6.4 RETARDATION PRESSURE Retardation pressure relates to a closed system with a low level of water filling. The level of water filling usually ranges from 0 to 50% of the total volume of the tunnel. The purpose of the water filling is to ensure favourable deposition of the blasted rock. It is very important to avoid the rock being transported to the gate. A computer program to calculate the pressure has been developed by The Norwegian Hydrodynamics Laboratories (NHL-report no. 3, project no. 601143). The program is based on iteration of the water inflow. For each time step, the energy balance is found in such a way that the energy conveyed to the water masses in the tunnel is equal to the energy used by friction, to compress the air and to accelerate the water. The friction is modelled as a singular loss characterised by a singular loss coefficient C, and a friction of the tunnel wall expressed as Manning's Number M. Figure 6.3 shows curves for the maximum pressure Hm, dependent on the static pressure H0 and the length of the tunnel L. The tunnel has a small cross section, A = 10 m2, and no water filling.

6. HYDRODYNAMICS 6.4 Retardation Pressure

55

��

Figure 6.3 Maximum retardation pressure as a function of static pressure and

length of the tunnel. Tunnel cross-section 10 m2 and no water filling.

6. HYDRODYNAMICS 6.5 Upsurge in the Gate Shaft

56

6.5 UPSURGE IN THE GATE SHAFT Maximum surge is normally limited to the floor of the gatehouse. The gates are usually installed at the time the piercing is performed. The basis for the calculation model is the difference in water level between the gate shaft and the reservoir. Without loss of head between the reservoir and the gate shaft, the upsurge would be as high above the reservoir level as it was below the reservoir level before the blasting of the plug. The upsurge in the shaft will normally be in the range of 0.8 - 0.9 times the full upsurge. Detailed estimations can be made based on the following assumptions: • = The blasting is regarded as a momentary removal of the plug. • = The effect of the blasting itself and the evacuation of the explosives gas and the

air pocket through the lake tap is neglected (i.e. free inflow of water through the lake tap at the time T = 0).

• = The head loss is said to be a singular loss. The cross section of the lower part of the gate shaft is usually very small in proportion to the rest of the shaft. The major part of the head loss will occur in this cross section. The other losses are neglected.

• = The calculations are performed as iterations. An example is shown in Appendix F5.

6. HYDRODYNAMICS 6.6 Transport of Mass

57

6.6 TRANSPORT OF MASS

6.60 Introduction The blast pit directly beneath the plug is designed to catch the rock that is blasted from the plug. The efficiency of the pit varies, depending on the type of piercing. Still, the pit should be excavated to catch sediments that will be transported through the plug during the operation of the power plant. The design, volume and placing of the pit cannot be calculated with certainty in advance, but laboratory model testing may give important information.

6.61 Open System Because of the water filling, the inflow velocity will be low. Hence, it is assumed that the blasted rock will fall down by its own weight and the pit has to be placed according to this.

6.62 Closed System The inflow velocities right after the blasting is very high. Thus, any trapping of the mass will be very difficult. Some of the mass will be caught in the trap, but the rest must be transported along the tunnel. Blocks of rock must not reach the gate. The water filling makes the catching of mass easier. If the gate is totally covered by water, it will be better protected against damage. It is very important to observe that the water filling reduces the head losses of the water inflow. Hence, the maximum pressure against the gate will increase.

6. HYDRODYNAMICS 6.7 Model Testing

58

6.7 MODEL TESTING

6.71 Open System Standard problems to be analysed are: • = Will the gate construction withstand the maximum pressure during the piercing? • = Will the upsurge in the shaft cause any damage to the installations in the shaft or

to the gatehouse? • = Will the blasted rock damage the gate construction or prevent the closing of the

gate? • = Is the rock pit correctly positioned and is the volume large enough? • = What is the best level of water filling in the tunnel and the shaft? • = Will the air in the air pocket evacuate through the plug? • = Does the shaft need a temporary contraction to reduce the maximum upsurge in

the shaft? The conditions of open system piercing are relatively easy to determine and the problems will be solved in most cases. Where there is a very complicated tunnel system and geometry, it may be necessary to carry out model testing in the laboratory.

6.72 Closed System While planning a complicated lake tap, it may be necessary to use model testing. Lack of detailed knowledge concerning the head losses for the water inflow makes the theoretical calculations somewhat uncertain.

REFERENCES

59

1. NIF-course no. 7241 Tunnelutslag under vann - terskelsprengning,

Kursdagene NTH, januar 1986.

(Underwater tunnel piercing and threshold blasting. Industrial Seminar at the

Norwegian Institute of Technology, January 1986. Papers in Norwegian.)

Paper 1 Chr. F. Grøner: Introduction.

Paper 2 John Johansen: Underwater tunnel piercing.

Paper 3 Bjørn Buen: Preinvestigations - Choice of piercing site.

Paper 4 Marvin Aarvoll: Design of tunnel system and piercing type.

Paper 5 John Johansen: Detailed planning of the blasting.

Paper 6 Øivind Solvik: Types of piercing - Necessary analysis of problems.

Paper 7 Halvor Kjørholt: Use of numerical modelling.

Paper 8 Jann A. Sandvik: Use of laboratory modelling.

Paper 10 Jann A. Sandvik: Measuring instruments - Control procedures for

tunnel piercing.

Paper 11 Terje Holm Nygård: Choice of construction equipment and

construction method.

Paper Øystein Stormo: Piercing from a 35 m2 tunnel at 85 m water depth,

Lake Ringedalsvatn.

REFERENCES

60

Paper Marvin Aarvoll: The Vestfjorden Sewage System (VEAS) - Central

Plant West. Piercing from the discharge tunnel.

2. Rapport fra Vassdragsregulantenes forening: Tunnelutslag under vann. En

oppsummering av eksisterende erfaring. Asker, april 1986.

(Report from Vassdragsregulantenes forening: Underwater Tunnel Piercing. A

summary of existing experience. Asker, April 1986.)

3. Norsk Hydro, reports by Arne B. Vagstein:

a) Utslag Styggevatn/tilløpstunnel Jostedal kraftverk, april 1987. (Piercing of

Lake Styggevatn/headrace tunnel of Jostedal Hydropower Plant, April 1987.)

b) Utslag Andrevatn og Tredjevatn, Ulla-Førre anleggene, april 1985.

(Piercing at Lake Andrevatn and Lake Tredjevatn, Ulla-Førre Hydropower

Plant, April 1985.)

c) Utslag Undeknutvatn, Ulla-Førre anleggene, januar 1985. (Piercing at Lake

Undeknutvatn, Ulla-Førre Hydropower plant, January 1985.)

4. Selmer Furuholmen: 47 utslagssprengninger utført fra 1950 til 1986, mai

1986. (47 piercing performed from 1950 to 1986, May 1986.)

5. Project report by Aud Røsbjørgen and Jarle Øverland: Tunnelutslag under

vann, videregående kurs i Anleggsteknikk, Institutt for anleggsdrift, NTH,

våren 1987.

(Underwater Tunnel Piercing, advanced course in Construction Engineering,

The Norwegian Institute of Technology, May 1987.)

6. Several reports from The Norwegian Hydrodynamic Laboratories, SINTEF,

Trondheim, concerning underwater tunnel piercing. 1970 - 1987.

REFERENCES

61

7. Papers presented at The Rock Blasting Conference in Oslo, arranged by The

Norwegian Tunnelling Society:

1965 Ragnar Hegstad: Orientering om forsøk med utslagssprengning.

(Experiments with tunnel piercing.)

1965 Chr. F. Grøner: Orientering om forsøk med skytegroper. (Experiments

with rock pits.)

1966 Ludvig Baumann: Tunnelutslag under vann og sprengning av terskler.

(Underwater tunnel piercing and threshold blasting.)

1980 Øystein Stormo: Utslag for 35 m2 tunnel på 85 m dyp, Ringedalsvatn,

Oksla kraftanlegg, Tyssedal. (Piercing from a 35 m2 tunnel at 85 m water

depth, Lake Ringedalsvatn, Oksla Hydropower Plant, Tyssedal.)

1983 Øivind Solvik: Tunnelutslag under vann. Prosjekteringskriterier,

risikomomenter og praktiske utforminger. (Underwater tunnel piercing.

Design criteria, risk factors and practical design.)

1984 Per Nordsletten and Gunnar Asting: Gjennomslag til

undervannskonstruksjoner, gassterminalen Kårstø.

(Piercing into underwater structures at the Kårstø Gas Terminal.)

1986 Rolf Medalen Krogh: Utslagssalver i Blåsjømagasinet, Ulla-Førre

anleggene. (Tunnel piercing at the Blåsjø Reservoir, Ulla-Førre Hydropower

Plant.)

1986 Sigmund Hoel: Utslag i sjøen for ilandføring av oljeledning ved Sture i

Øygarden. (Tunnel piercing into the sea for shore approach of the oil pipeline

at the Sture Terminal, Øygarden.)

REFERENCES

62

1987 Sigmund Hoel: Utslag i sjøen for ilandføring av oljeledning ved Sture i

Øygarden. (Tunnel piercing into the sea for shore approach of the oil pipeline

at the Sture Terminal, Øygarden.)

1987 Erik Dahl Johansen: Utslag Fossvatn, Kobbelv kraftanlegg. (Piercing of

Lake Fossvatn, Kobbelv Hydropower Plant.)

1994 Per Aftreth, Finn Hvoslef: Verdens dypeste utslag under vann ( The

deepest subsea piercing of the world)

8. Dyno Konsulent A.S, Oslo: Report P-319, Tunnelutslag i vann (Underwater

tunnel piercing):

Phase 1: September 1980

Phase 2: May 1981

Phase 3: January 1985

9. Frode Færøyvik: Fra feisel til fullprofil - Fjellsprengere i samfunnets tjeneste.

Norsk Forening for Fjellsprengningsteknikk, 1989. (From hand hammers to

tunnel boring machines - History of tunnelling in Norway. The Norwegian

Tunnelling Society, Oslo 1989.)

10. Øivind Solvik: Underwater piercing of a tunnel. Water Power & Dam

Construction, November 1984.

11. B. Berdal, B. Buen, J. Johansen: Lake tap - The Norwegian method.

Tunnelling '85, paper 9.

REFERENCES

63

12. Kåre Rokoengen: Hjelpemidler ved grunnundersøkelser på

kontinentalsokkelen. IKU-publikasjon nr. 80, oktober 1976. (Methods of

preinvestigations of the continental shelf. IKU publication no. 80, IKU,

Trondheim October 1976.)

13. O. Johannessen, J. Skutle, B. Thorsen: Project Report 5-83 TUNNELLING

Prognosis. The Division of Construction Engineering, The Norwegian Institute

of Technology, 1983. (Updated version: Project Report 2-88, see page 175.)

14. O. Johannessen, J. Skutle, B. Thorsen: Project Report 6-83 TUNNELLING

Costs. The Division of Construction Engineering, The Norwegian Institute of

Technology, 1983. (Updated version: Project Report 3-88, see page 175.)

APPENDIX A. Previous Edition of the Report

64

A. PREVIOUS EDITIONS OF THE REPORT The first version of the project started in 1987 when a working group was formed at the Division of Construction Engineering at The Norwegian Institute of Technology at the University of Trondheim. The group comprised faculty and students at the division: Amund Bruland Odd Johannessen Karsten Myrvold Aud Røsbjørgen Jarle Øverland The project work was followed by a diploma thesis by Aud Røsbjørgen and Jarle Øverland, who also finalised the Project Report 19-88 UNDERWATER TUNNEL PIERCING. The project report was completed in the spring of 1991, in a Norwegian and an English edition. This second edition of the report is edited by Steinar Roald and Gry Helle Nakstad.

APPENDIX B. Research Partners

65

B. RESEARCH PARTNERS The following external research partners have supported the project:

• =Statkraft anlegg as • =Norwegian Public Roads Administration • =Statsbygg • =Scandinavian Rock Group AS • =NCC Eeg-Henriksen Anlegg AS • =Veidekke ASA • =Andersen Mek. Verksted AS • =DYNO Nobel • =Atlas Copco Rock Drills AB • =Tamrock OY • =The Research Council of Norway

APPENDIX C. Definitions and Parameters

66

C. DEFINITIONS AND PARAMETERS

Expression Definition

NWL Natural water level 4

WL Actual water level 7

HWL Highest regulated water level 7

LWL Lowest regulated water level 7

seismic Caused by earthquake or artificial earth vibrations 14

refraction Deflection from a straight path 14

ROV Remote Operated Vehicle 17

pulse Shock wave of brief duration 25

confinement The term constriction and fixation is also used.

For a complete description, refer to U. Langefors and

B. Kihlström: The Modern Technique of Rock Blasting,

Stockholm 1963, 42

Group 1, 2, 3 refers to classifying of detonators according to current

Norwegian Government regulations 44

mWc metre water column (= 0.1 bar) 49

APPENDIX D. Basis of the Report

67

D. BASIS OF THE REPORT Information from the following sources has been used during the preparation of the project report. We would like to acknowledge and thank the companies and institutions for their positive support: • = Veidekke ASA (Aker Entreprenør) • = Selmer ASA (Selmer Furuholmen A/S) • = A/S Geoteam • = Norconsult (Berdal Strømme A.S) • = The Department of Geology and Mineral Engineering, NTNU Trondheim • = Dyno Consult A/S • = IKU - The Continental Shelf and Petroleum Technology Research Institute • = SINTEF NHL - Norwegian Hydrotechnical Laboratories • = NIF-course no. 7241, January 1986, Underwater Tunnel Piercings • = Norsk Hydro ASA • = Noteby A/S • = Statkraft (The Norwegian State Power Board) • = ENFO (Vassdragsregulantenes forening) • = Rescon A/S • = The Swedish Geological Company Ltd. • = Trondheim Mørtelverk A/S • = The Rock Blasting Conference (Fjellsprengningskonferansen) in Oslo

1963 – 1999, several papers

APPENDIX E. Some recent, Large Underwater Piercings in Norway

68

E. SOME RECENT, LARGE UNDERWATER PIERCINGS IN NORWAY

Water depth refers to the level of the tunnel plug.

1. Kobbelv Hydropower Plant: 7 piercings, with Lake Fossvatn at 100 m water

depth, 1987, [7].

2. Ulla-Førre Hydropower Plant: 5 piercings at the Blåsjø Reservoir, with 101 m

water depth as a maximum, 1985 and 1986, [7].

3. Oksla Hydropower Plant, Tyssedal: Piercing from the headrace tunnel into Lake

Ringedalsvatn, with 85 m water depth, 1980, [1] and [7].

4. Eidfjord Hydropower Plant: Piercing from the headrace tunnel into Lake

Rembesdalsvatn, with 25 m water depth, 1977, [8].

5. Lomi Hydropower Plant: Piercing at Lake Lomivatn, with 75 m water depth,

1978, [8].

6. The Kårstø Gas Terminal: 2 piercings into underwater structures, with 10 m and

30 m water depth, 1984, [7].

7. The Sture Terminal: Piercing into the sea at 80 m depth. Tunnel for shore

approach of an oil pipeline from the North Sea Oseberg Field to Sture,

Øygarden, 1986, [7].

8. The Troll Field: The shore approach for the gas pipelines is a subsea tunnel

system with comprises tunnels, shafts, concrete works and installation of steel

support cones for rise bundels, and 3 piercings at 160 m depth (1994), [7].

APPENDIX F1. Seismic Methods

69

F1. SEISMIC METHODS

F1.1 Reflection Seismics The principle of this method is similar to acoustic sounding. However, as an echo sounder only gives the profile of the bottom of the lake, reflection seismic uses more powerful pulses with lower frequency, which also makes it possible to determine the boundaries between deeper strata. The equipment for reflection seismic consists of a source of pulses (transmitter) and hydrophones (receiver). The equipment is mounted on a cable which is towed by a boat, see Figure F.1. The position of the boat is known exactly by means of radio transmitters (beacons) on shore. The most usual transmitters in reflection seismics are the Boomer and Sparker. The Boomer utilises a mechanical pulse source and is able to register strata thickness of 2 m and more. The Sparker utilises an implosion pulse and has better penetration capacity than the Boomer, but the resolution is lower, see Table F.1.


70

Figure F.1 Principle of reflection seismics measurement and registration.

[12, Figure 12]

Method Penetration Depth, m Resolution, m Echo sounder Low -

Boomer 50-200 Approx. 2 Sparker 100-1000 >2

Table F.1 Specifications of the seismic methods. The pulses are reflected from the strata limits and the thickness of each stratum can be determined.


71

In addition to the thickness of the strata, the material parameters can be decided on the basis of the seismic velocity in each stratum. It is important to be aware of the fact that the hydrophones do not only receive the reflected waves, but also the direct wave through the water, and the multiples. In deep, narrow lakes and fjords, the reflections from the sides may arrive earlier than the reflections from the bottom. This is called side reflection, see Figure F.2. The problem is avoided by towing the equipment near the bottom. Because reflection seismic is a quick and comparatively inexpensive method, it is used for initial mapping of larger areas. This facilitates the evaluation of possible sites for the underwater piercing, and the most promising may be examined more closely by more accurate methods.

Figure F.2 Side reflection. Costs The cost of reflection seismic is about NOK 1500 - 2000 per kilometre of profile, excluding rigging costs of approximately NOK 100 000.

F1.2 Refraction Seismic With refraction seismic, the acoustic waves will be reflected and refracted similarly to reflection seismic, see Figure F.3.


72

Figure F.3 Principle of refraction seismic.

[1, paper 3] When a shock wave reaches the limit between two strata, one part of the energy will be reflected, one part will propagate along the limit and one part will pass through the limit. Placing hydrophones in known positions at the bottom of the lake enables the measurement of the travel time for known distances. The seismic velocity of the stratum may be calculated and measurements are registered directly in a diagram of travel time, see Figure F.4.


73

Figure F.4 Diagram of travel time. The best result is obtained by measuring on the bottom. The hydrophones are connected to a cable that also transfers the signals. The cable is placed from a boat, and if necessary, the cable is fastened on shore, see Figure F.5. The position of the cable is determined by measuring the position of the markers. The position may be found indirectly by measuring the travel time through the water from a shot point. Accurate positioning is important to achieve satisfactory results. The method has been used at depths of less than 400 m.


74

It is also possible to do refraction seismic measurements from a boat or from the ice on frozen lakes, but these measurements are not as precise as measurements made on the lake bottom. Normally, explosives are used to generate shock waves.

Figure F.5 Example of refraction seismic measurement. The seismic velocity in the different strata indicates what type of material the strata consist of and the quality of the material, e.g. fractured rock. As shown in Figure F.6, there is some overlap in seismic velocities for the different geological materials. A refraction seismic measurement may result in several possibilities concerning the type of material in a stratum. Hence, it is also necessary to evaluate other geological data.


75

��

�� !��

��

��

�� "�� !��

� ��

��

Figure F.6 Seismic velocities in sediments and rock. Refraction seismic measurements are only of value if they are carried out and interpreted by qualified personnel who know the limitations of the method. Costs The time and cost requirements depend on the water depth, the weather and the current conditions. Usually the costs are between NOK 100 to NOK 500 per m of profile, including rigging costs.

APPENDIX F2. Consumption of Grouting Media

76

F2. CONSUMPTION OF GROUTING MEDIA IN SYSTEMATICALLY FRACTURED ROCK MASS.

The model of calculation is idealised. It is presupposed that the joints are open and continuous. However, the joints are often filled with water, clay or minerals. Besides, all the joints will hardly be filled with grouting media. This will make the calculated consumption larger than the actual consumption. The following assumptions are made: • = The round length is 4 m. • = Grouting is performed every third round. • = The length of the grouting holes is 19.3 m and the overlap is one round,

see Figure F.7. • = The grouting covers approximately 5 m outside the contour around the cross

section, see Figure F.8. • = The holes are placed at equal intervals around the contour of the tunnel and

drilled as an outward fan.

Figure F.7 Grouting screen.


77

With a grouting and tunnel cross section as shown in Figure F.8, the grouted volume per metre of tunnel is: V1 = 5 ⋅ 4 + 0.5 ⋅ π ⋅ 2.52 = 29.8 m3/m V2 = 0.5 ⋅ π ⋅ (7.52 - 2.52) = 78.5 m3/m V3 = 2 ⋅ 5 ⋅ 4 = 40.0 m3/m V4 = 5 ⋅ 5 = 25.0 m3/m V5 = 2 ⋅ 0.25 ⋅ π ⋅ 52 = 39.3 m3/m

Vinj = 212.6 m3/m

��

��

��

��

��

��

Figure F.8 Grouting and tunnel cross section.


78

��

��

��

Figure F.9 Placing of the grouting holes.

Joint Class Distance between

joints [cm] Amount of

joints [m2/m3] 0 - -

0 – I 160 0.6 I 40 2.5 II 20 5.0 III 10 10.0 IV 5 20.0

Table F.2 Joint classification. From Project Report 1D-98 HARD ROCK TUNNEL

BORING Geology and Preinvestigations.

Joint Class Joint Opening

[mm] 0 - I I II III IV

0.5 0.0003 0.0013 0.0025 0.005 0.01 1.0 0.0006 0.0025 0.005 0.01 0.02 1.5 0.0009 0.0038 0.0075 0.015 0.03 2.0 0.0012 0.005 0.01 0.02 0.04

Table F.3 Joint volume.


79

Joint Class Joint

Opening [mm] 0 – I I II III IV

0.5 0.064 0.28 0.53 1.06 2.13 1.0 0.13 0.53 1.06 2.13 4.25 1.5 0.19 0.81 1.59 3.19 6.38 2.0 0.26 1.06 2.13 4.25 8.50

Table F.4 Joint volume per tunnel metre [m3/m]. Example Assuming that the joint opening is 1.5 - 2.0 mm, the use of cement-based suspensions is possible. For a mix ratio of water and cement of 1: 1.5, 0.75 kg cement gives 1 litre of grouting solution (information from Rescon's data sheet of Cemsil).

Joint Class Joint Opening

[mm] 0 – I I II III IV

1.5 143 608 1193 2393 4785 2.0 195 795 1598 3188 6375

Table F.5 Consumption of cement [kg] per tunnel metre.

APPENDIX F3. Blasting with Short Rounds

80

F3. BLASTING WITH SHORT ROUNDS When approaching the breakthrough area, the advance per round must be reduced. Referring to the project reports 2B-95 TUNNELLING Prognosis for Drill and Blast and 2C-95 TUNNELLING Costs for Drill and Blast, it is possible to find the time consumption and costs when the advance per round is reduced down to 1 m. The data in the reports are based on a standard round length of 5.0 m. For round length different from 5.0 m, the values must be corrected.

F3.1 Round Cycle Time Estimation of the round cycle time for a tunnel with the following data: Cross section 30 m2 Rock blastability medium Rock drillability DRI = 49 (medium) Rock drills 3 AC COP 1838 Drillhole diameter 45 mm Large drillholes 102 mm Using Project Report 2B-95, the round cycle is divided into four major operations: I Drilling, charging, blasting II Ventilation III Loading and hauling IV Scaling The parameters and estimation modules used in Tables F.6 and F.7 are described in detail in the two project reports referred above.

APPEND

IX F3. Blasting w

ith Short Rounds

81

1

100 0.885

56

1

8.08 0.42 0.62

11.97

0.51

2.15 23.15

1.6

160 0.895

56

1

12.93 0.67 0.64

12.37

0.82

2.73 29.52

2.1

210 0.905

57

1

17.27 0.88 0.66

12.98

1.10

3.27 35.50

3

300 0.92 58

2

25.11 2.51 0.68

14.05

1.62

4.38 47.67

4

400 0.945

60

3

34.63 5.03 0.72

15.84

2.27

5.83 63.60

5 30

High Medium

45 63

500 1

63 102

3 AC-COP 1838

3 49

231 100 231 43

99.33 45.45 6.29 0.75

17.25 Medium 0.018

2.97 0.105

7.24 79.21

Unit

m m2

mm

cm

mm

cm/min %

cm/min %

cm/min min min min min

min

min min

Parameter

lb As

SPR d1 ab lb kbl ah dg ag

am DRI b45 kl bl kg bg th tg ff tf

fk tk fsa tsa Tb

Drilled round length Tunnel cross section Skill level Blastability Drillhole diameter Number of drillholes basic round length Drilled length Correction drilled length Number of holes exclusive large drillholes Diameter large drillholes Number of large drillholes Type of drilling hammers Number of drilling hammers Drillability Penetration rate 45 mm drillhole Correction penetration rate for d1 in relation to 45 mm Penetration rate d1 Correction penetration rate for dg in relation to 45 mm Penetration rate dg

Drilling time Drilling time large drillholes Time for moving per hole Time for moving Rock wear properties Bit changing factor Time for bit changing Simultaneousness Time due to simultaneousness Necessary drilling time

Table F.6 Time consum

ption per round cycle and weekly advance rate.

APPEND

IX F3. Blasting w

ith Short Rounds

82

1 23.15

31.5

8.18 81.8

33.102 25.46

4.21 42.2

171 32

1.6 29.52

32

8.94 89.5

52.9632 40.74

5.91 59.2

196 45

2.1 30.50

32.5

9.66 96.7

69.5142 53.47

7.32 73.3

217 53

3 47.67

33

11.06 110.7

99.306 76.39

9.87 98.8

256 64

4 63.60

33.5

12.89 129.0

132.408 101.85

12.69 127.0

303 72

5 79.21

3 35.5

19 14.84 148.6

17 Cat 980 F

Truck - Load and haul with 110 m between the niches 78

1.226 90

165.51 127.32

12.5 15.52 155.3

30 351 78

Unit m

min

min min min min min

asm3/h

% asm3 min min min min min min

m/week

Parameter lb Tb

Tl Trb Ttb

l fo br v Tu Trl Ttl

Drilled round length Necessary drilling time Number of charging lines Determining charging time Rig time drilling, charging, blasting Incidental lost time drilling, charging, blasting Sum for drilling, charging, blasting Ventilation break Type of loader Transport equipment Gross loading capacity Factor of overbreak Advance per round Actual volume per round Loading time per round Rig time loading and hauling Incidental lost time loading and hauling Sum loading Scaling time, scaling from jumbo TOTAL TIME PER ROUND CYCLE Weekly advance rate

Table F.6 Time consum

ption per round cycle and weekly advance rate.


83

� � �

� � �

� � �

� � �

� � �

�

� � �

� � � �

�

Figure F.10 Round cycle time as a function of drilled round length.

� �

� �

�

� �

�

� �

� �

�

� � � � �

� � � �

�

Figure F.11 Weekly advance rate as a function of drilled round length.


84

F3.2 Excavation Costs The assumptions are as shown in section F3.1. Project Report 2C-95 divides the tunnel costs into the following categories: I DRILLING, CHARGING, BLASTING AND SCALING II LOADING III HAULING IV ADDITIONAL WORK The price level in this section is January 1995. Estimated cost index from January 1995 to January 2000 is 1.19.

APPEND

IX F3. Blasting w

ith Short Rounds

85

1

1.23

2460

1.595

1945.9 7088.9

7798

1.6

1.155

2310

1.54

1878.8 6871.8

7559

2.1

1.12

2240

1.43

1744.6 6667.6

7334

3

1.055

2110

1.27

1549.4 6342.4

6977

4

0.995

1990

1.095

1335.9 6008.9

6610

5 30 3

Trackless tunnelling 49

Medium 1 45

2000 1

1 5 1

2000 Cat 980 F Truck - Load and haul with 110 m between the niches

500 1.0 500

Trackless tunnelling 1200

1 1200 154 75

1429 335 157 145 117 754 1220

1 1

1220 5903

10 1

6493

Unit m m2 km

mm NOK/m

%

NOK/m

NOK/m

NOK/m

NOK/m

NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

NOK/m NOK/m

%

NOK/m

Parameter

As

DRI

kT

cd0 kbl kle

kdn cd

cl0 kle cl

ch0 kut ch cr ctip ct cv ce cw cm ca cla0 kbl kll cla cs0

kp cs

Drilled round length Tunnel cross section Tunnel length Excavation method Drillability Blastability Adit: Horizontal Drillhole diameter Total costs drilling, charging, blasting… Correction for drilled length Correction for tunnel length and job training effect Proportion of dynamite Correction for dynamite proportion Corrected cost drilling, charging, blasting.. Loading equipment Loading cost Correction for tunnel length and job training effect Corrected loading costs Type of transport Hauling costs Correction for utilisation Corrected hauling costs Costs for road pavement Tip costs Total hauling costs Ventilation Electrical installations Water supply Miscellaneous costs Additional costs Labour Correction for drilled length Correction for tunnel length Corrected labour costs Sum elemental costs 10 % for uncertain assumptions Correction for price level TOTAL STANDARD COST excl. rock support

Table F.7 Total standard cost.


86

� � �

� � �

� � � �

� � � �

� � �

� �

� �

�

� � � �

� � � �

Figure F.12 Excavation costs as a function of drilled round length. Price level 1995.

APPENDIX F4. Example of Application for Charging Plan

87

F4. EXAMPLE OF APPLICATION FOR CHARGING PLAN

F4.0 Introduction The design of a drilling pattern and a plan for charging and coupling is illustrated by the planning of the piercing round at Lake Styggevatn of the Jostedal Hydropower Project in Norway, an open system piercing. General information: • = At water level 1175 m, the piercing will be performed at a depth of about 70 m • = The cross section of the plug is 27 m2. • = The thickness of the plug is 6.5 m, which is determined by probe drilling into

the water. • = The drillhole diameter is 45 mm.

��

��

Figure F.13 An outline of the piercing at Lake Styggevatn. See also Figures 1.1

and 6.3.


88

F4.1 Specific Charging Specific charging is found from Section 5.2. The standard consumption of explosives for a tunnel with a cross section of 27 m2, is approximately 1.7 kg/sm3. Basic specific charging 3.4 kg/sm3 Figure 5.1 Additional charging for 70 m water depth 0.01 · 70

0.7 kg/sm3

Section 5.2

Sum 4.1 kg/sm3 10 % addition for lost or unusable holes 0.1 · 4.1 0.4 kg/sm3 Total specific charging 4.5 kg/sm3

By using the diagram in Figure 5.1, specific charging is found to be 4.6 kg/sm3. In the following calculations, a specific charging of 4.6 kg/sm3 is used.

F4.2 Necessary Number of Drillholes The following explosives are used, Section 5.2: • = Extra dynamite (35 · 600 mm cartridges). • = Plastic cartridges. • = The amount of explosives in one cartridge is 0.95 kg. Explosives data refer to explosives manufactured by DYNO Nobel in Norway. The following is assumed: • = No compression of the explosives in the drillhole. • = Uncharged length is 0.5 m. • = 0.5 m of rock at the end of each drillhole. Charged length of one drillhole is: 6.5 m - (2 ⋅ 0.5 m) = 5.5 m Number of cartridges in one drillhole: 5.5 m / 0.6 m ≈ 9 Amount of explosives in one drillhole: 0.95 kg ⋅ 9 = 8.5 kg Necessary number of drillholes in the plug: (4.6 ⋅ 27 ⋅ 6.5) / 8.5 = 95


89

Specific charging 4.6 kg/sm3 Cross section 27 m2 Thickness of the plug 6.5 m Explosives in each hole 8.5 kg

F4.3 The Cut In drill and blast tunnelling with a cut of parallel holes, it must be possible for the blasted rock to expand 80% to ensure the throw of the blasted rock. Piercing rounds have high specific charging which develops a lot of explosion gases. Then there are two throw directions. Both these facts result in good throw of the blasted rock even though the possibilities of expansion are less than 80%. The reason for reducing the possibility to expand by increasing the distance between the large holes is to reduce the risk of intersection of the large holes. Figure 5.3 shows that the necessary total area of large holes is 900 cm2. The following cut is chosen to achieve a rectangular cut that may be placed anywhere in the drilling pattern: 12 holes of 102 mm diameter give 980 cm2 of large hole area.

F4.4 Specific Drilling Charged holes 95 holes of 6 m = 570 m Large holes 12 holes of 6 m = 72 m Specific drilling (570 + 72)/(27 · 6.5) = 3.66 dm/sm3 The diagram in Figure 5.2 gives 3.85 dm/sm3


90

F4.5 Drilling Pattern As mentioned in Section 5.1, the cross section should be nearly rectangular to make it possible to drill systematically with an even distance between the holes. The corners are broken to reduce the confinement of the corner holes. For details of the drilling pattern, see Figure F.14. • = The distance between the holes is chosen to be 60 cm. • = For the cut, the distance is reduced to 30 cm • = The cut is placed after the drilling of the charged holes is finished. It is then

possible to place the cut in the most favourable position with regard to rock quality and confinement of holes

Charged holes 96 holes of 6 m = 576 m Large holes 12 holes of 6 m = 72 m Specific drilling (576 + 72)/(27 ⋅ 6.5) = 3.7 dm/sm3


91

Figure F.14 Drilling pattern (45 mm charged holes and 102 mm large holes).

Dimensions in cm.


92

F4.6 Charging Plan The chosen drilling pattern results in specific charging of: (96 ⋅ 8.5)/(27 ⋅ 6.5) = 4.6 kg/sm3 96 holes are to be charged and two detonators are placed in each hole, see Section 5.2. Total number of detonators in the piercing round is 96 ⋅ 2 = 192 The two detonators in each hole are coupled in two separate series, which then are coupled in parallel, see Figure F.15.

Figure F.15 Firing plan with detonator numbers. There are two detonators in

each charged hole.


93

F4.7 Coupling Plan For the piercing round, (German) millisecond detonators (Group 1) with extra insulation of the cables are used, see Section 5.2. These detonators are available with 6 m or 10 m cables. For Lake Styggevatn, the length of the drillholes was 6 m. To ensure sufficient length of the cables, the 10 m option was chosen. The following specifications are taken from product data sheets. Resistance per detonator (10 m cable): Rd = 2.4 ohm Initiating pulse: 5 mJ/ohm Lowest current of initiation: 0.28 A Critical current: 1.1 A The critical current is the current which ensures 99.9 % probability of initiating detonators coupled in series. Resistance per series: Rs = 2.4 ohm ⋅ 96 = 230.4 ohm Series-parallel resistance: 1/Rsp = 1/Rs1 + 1/Rs2 Rsp = 230.4/2 = 115.2 ohm Connecting cable: 2 · 2.5 mm2 Resistance per 100 m of cable: 1.4 ohm Length to the place of firing: 730 m Resistance in the connecting cable: Rk = 1.4 ⋅ 7.3 = 10.3 ohm Total resistance at the place of firing: Rtot = Rsp + Rk = 115.2 + 10.3 = 125.5 ohm The blaster is chosen in accordance with Section 5.4. The recharging voltage for the chosen instrument is: 1500 V The loss of voltage in the wire is about: 100 V


94

Ohms law: U = R ⋅ I U = voltage R = resistance I = current Current: I = U/R = 1400/125.5 = 11.2 A Branch current: Ib = 11.2/2 = 5.6 A With a safety factor of 1.5, the branch current will be: 5.6/1.5 = 3.7 A > 1.1 A The coupling plan is shown in Figure F.16.

Figure F.16 Coupling plan (two parallel series of 96 detonators each).

APPENDIX F5. Example of Application for Upsurge in the Gate Shaft

95

F5. EXAMPLE OF APPLICATION FOR UPSURGE IN THE GATE SHAFT Parameters used in the calculation: δH = the difference in water level between the gate shaft and the reservoir δT = the length of one time step (iteration step) ΣL/F = the sum of length divided by the cross section of each part of the string of water involved Qn = the flow of water in the tunnel at the end of time step n Qn-1 = the flow of water in the tunnel at the end of time step n-1 Hn = the water level in the gate shaft at the end of time step n Hn-1 = the water level in the gate shaft at the end of time step n-1 δQ = change in the flow of water during time step n Q = average flow of water in the tunnel during time step n Fi = cross section of the narrow pass at the bottom of the shaft Fs = cross section of the gate shaft g = acceleration of gravity = 9.81 m/s2 C = singular loss coefficient for the narrow pass expressed as velocity

head. Primarily, the coefficient is dependent on the shape of the narrow pass, but the value of the coefficient will always be near 1.0. In most cases, the value 1.0 is used.

To get satisfactory precision, the time step δT should be long enough to allow at least 10 iteration steps before achieving the maximum upsurge. However, there is not much to gain by increasing the number of steps to 20. The equation of the upsurge:

Tg

FQFQCH

FLgQ ii δδδ ⋅

⋅⋅

⋅−⋅Σ

= )2

//(

/ [F.1]

QQQ nn δ+= −1 [F.2]


96

The basis of the calculations is the situation shown in Figure F.17. Furthermore, the following is assumed: Singular loss coefficient for the narrow pass: C = 1.0 Cross section of the gate shaft: Fs = 2.5 ⋅ 2.5 = 6.25 m2 Cross section of the narrow pass: Fi = 1.0 ⋅ 2.5 = 2.5 m2 Total length divided by cross section area: ΣL/F = (434/27) + (60/6.25) = 25.6 m-1 Constant part of equation [F.3]: g/(ΣL/F) = 9.8/25.6 = 0.38 The calculation starts with a difference in water level between the gate shaft and the reservoir of 15 m, i.e. δH = 15.

n δT g/(ΣL/F) g

FQFQC ii

⋅⋅

⋅2

// δQ Qn Q δH

0 0 0 15 1 1 0.38 0 5.7 5.7 2.9 14.5 2 1 0.38 0.3 5.4 11.1 8.4 13.2 3 1 0.38 1.0 4.6 15.7 13.4 11.1 4 1 0.38 2.0 3.5 19.2 17.5 8.3 5 1 0.38 3.0 2.0 21.2 20.2 5.1 6 1 0.38 3.7 0.5 21.7 21.5 1.7 7 1 0.38 3.8 -0.8 20.9 21.3 -1.7 8 1 0.38 3.6 -2.0 18.9 19.9 -4.9 9 1 0.38 2.9 -3.0 15.9 17.4 -7.7 10 1 0.38 2.1 -3.7 12.2 14.1 -10.011 1 0.38 1.2 -4.3 7.9 10.1 -11.612 1 0.38 0.5 -4.6 3.3 5.6 -12.513 0.4 0.38 0.1 -1.9 1.4 2.4 -12.714 0.3 0.38 0 -1.4 0 0.7 -12.7

21−+= nn QQQ [F.3]

TFQHH

sgn δδδ ⋅−= [F.4]


97

According to the calculations, the maximum upsurge will be approximately 12.7 m above the water level in the reservoir (1187.7 metres above sea level). The floor of the gate house should be situated at a higher level.

��

��

��

��

�

�

��

��

��

� � � � � � � ��

� ��

��

Figure F.17 Upsurge of the piercing at Lake Styggevatn according to the

estimations.

APPENDIX F6. Example of Application for Explosives Gas Pressure

98

F6. EXAMPLE OF APPLICATION FOR EXPLOSIVES GAS PRESSURE The basis of the example is the piercing round at Lake Styggevatn, see Figures F.3 and F.5. According to the example in Section 5.52, the amount of explosives is 8.5 kg per hole. With 96 charged holes, the total amount of explosives is 816 kg. In the following, 820 kg of explosives is assumed. According to Figure 6.3, the following parameters are known: Volume of air at the face: 4500 Nm3 Water level Lake Styggevatn: 1175 m Water level in the air pocket: 1099 m Water level in the gate shaft: 1160 m Level of the gate threshold: 1094 m Pressure against the gate before blasting:

Pressure in the air pocket before blasting:

The ratio amount of explosives to volume of air at the face, is an important parameter for the resulting explosives gas pressure. The amount of explosives is usually determined in advance, while the volume of air has to be adapted to ensure an acceptable pressure increase. In this case the ratio is:

The pressure increase in the air pocket at blasting of the plug, is:

6.70.110

109411601 =+−=p (bar, absolute) [F.5]

1.70.110

109911600 =+−=p (bar, absolute) [F.6]

18.04500820

][Nmair of Volume[kg] explosives ofAmount

3 ===µ [F.7]


99

pg = maximum pressure in the air pocket after detonation (absolute) p0 = pressure in the air pocket before detonation (absolute) pa = atmospheric pressure (normal pressure) (absolute) κ = adiabatic exponent. The ratio pg/p0 is shown in Figure F.18. For Lake Styggevatn, the ratio is: pg/p0 = 1.3 The maximum pressure at the plug after detonation is: pg = 1.3 · p0 = 1.3 ⋅ 7.1 = 9.2 bar (abs) The increase in pressure in the air pocket after detonation is:pg - p0 = 9.2 - 7.1 = 2.1 bar The design pressure for the gate is:pd = p1 + 2 · (pg - p0) = 7.6 + 2 ⋅ 2.1 = 11.8 bar (abs) The design difference in pressure at the gate is: 11.8 - 1.0 = 10.8 bar This corresponds to a water level of 108 m above the gate threshold.

κκ

µ

�

��

�

�

⋅⋅��

��

�

+= 8.01

/1

o

a

o

a

o

g

pp

pp

pp

κ = 1.4 [F.8]


100

Figure F.18 Diagram of explosives gas pressure. See also Figure F.19.

��

Figure F.19 An outline of the piercing at Lake Styggevatn. Open system piercing.

APPENDIX G1. Example: Underwater tunnel Piercing at Lake Fossvatn

101

G1. EXAMPLE: UNDERWATER TUNNEL PIERCING AT LAKE FOSSVATN,

KOBBELV HYDROPOWER PLANT

G1. Situation

The Kobbelv Hydropower Plant is situated in the county of Nordland, approximately

67°30' N and 80 km north-east of the city of Bodø.

G2. Main Data of the Power Plant

Installed power: 2 · 150 MW

Average head: 600 m

Average yearly production: 710 GWh

Construction period: 1982-1987

The Norwegian State Power Board (Statkraft) was responsible for design, planning

and construction.

G3. Underwater Tunnel Piercings at Kobbelv

Site Type of

piercing

Cross

section

tunnel (m2)

Water

depth plug

(m)

Water

depth

gate (m)

Cross

section

plug (m2)

Voulme

of plug

(sm3)

Dictance

from gate

(m)

Amount of

explosives

(kg)

Lievsejavre closed 18 40 70 10 50 550 350

Reinoksvatn closed 16 65 85 10 50 465 350

Fossvatn open 51 100 119 40 270 370 1100

Varravæjekajavre open 35 34 52 30 150 90 1100

Lagnvatn West open 18 62 90 10 50 450 350

Langvatn East open 18 77 90 10 50 340 350

Kobbvatn open 51 6 8 51 425 10/600 2000

Table G.1 Data Kobbelv Hydropower plant


102

Further documentation of the piercing at Lake Fossvatn will be given. Lake Fossvatn has been regulated 10 metres above and 90 metres below the natural water level.

All the piercings were planned and performed by The Norwegian State Power Board

in cooperation with The Norwegian Hydrotechnical Laboratories and Dyno

Konsulent A.S. Dyno Konsulent A.S was responsible for the drilling and charging

plans.

G4. Data for the Piercing at Lake Fossvatn

The piercing was based on a compressible air pocket under the piercing plug and

water filling of the gate shaft, in order to reduce the upsurge and the maximum

pressure against the gate. The piercing round was blasted with the main gate closed,

which was designed for a maximum pressure of 181 metres of water.

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Figure G.1 Longitudinal section of the piercing. Natural water level is 610.8 m.


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G4.1 Geology

The rock type was a homogeneous granitic gneiss, with seismic velocity of

approximately 5000 m/s. The rock mass contained some joints and fractures and

consequently, 15000 kg of cement were used for grouting. Nevertheless, the water

leakages were still comparatively large. When the drilling of the piercing round was

finished, the leakage was more than 4.5 m3/min.

The choice of tunnel route and piercing site was based on seismic investigations and

mapping by sounding from the ice during the winter. The preinvestigations showed

1 - 3 metres of sediments at the piercing site.

G4.2 The Tunnel under Lake Fossvatn

The tunnel cross section was 51 m2. Systematic probe drilling was carried out the last

100 metres of the tunnel. The probe drilling showed good agreement with the seismic

preinvestigations. For each round, a 20 metre long hole was drilled upwards and

forward from the tunnel face. When the tunnel got close to the piercing area, several

holes of this type were drilled from the roof and the walls of the tunnel.

The tunnel was excavated with full round length (4.3 m drilled) except for the last

three rounds.

The shaft under the piercing plug was excavated by small rounds. Then the piercing

plug was trimmed and the piercing round was drilled before the rock trap was

excavated. The rock trap had a volume of 600 sm3 and was excavated by five rounds.


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G4.3 Gate Shaft and Gates

The main gate is a rolling gate and the revision gate is a gliding gate.

In order to limit the upsurge in the gate shaft after the round was blasted, it was

necessary to reduce the shaft cross section from 16 m2 to approximately 2.9 m2 at the

bottom of the shaft.

The pressure against the gate was measured by an electronic pressure sensor.

G4.4 The Piercing Plug

The area was 40 m2 and the thickness was 6.5 - 7 m, making the plug volume

approximately 270 sm3. The water pressure against the plug was approximately 100

m. The thickness of sediments above the plug was 1 - 3 m.


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G4.5 Drilling Pattern

Figure G.2 Placing of drill holes, including the cut. Measures in cm.

The cut is a double parallel hole cut, consisting of 8 uncharged 102 mm diameter

large holes and 21 charged 45 mm diameter holes.

The total number of 45 mm diameter charged holes is 159, including the charged

holes in the cut.


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Figure G.3 The cut. Dimension in cm.

It is important to make the drilling pattern systematic.

All holes were drilled until 50 cm of rock was left between the bottom of the hole and

the lake.


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Figure G.4 Drilling of the plug, longitudinal section. Dimension in cm.

G4.6 Charging Plan

The piercing round was charged by Extra Dynamit (35 · 600 mm cartridges) in all

charged holes. The explosives and the detonators were designed for 100 m water

pressure for 72 hours. The charge density in the hole was approximately 1.4 kg/m.

The length of each hole was checked. The uncharged hole length was 40 cm, which

was stemmed with expanded polystyrene and a wooden plug with a slot for the

initiating cables.

A 15 m high scaffolding was built from the bottom of the rock pit for the charging.

A small boat was used for access from the gate shaft to the plug during charging.


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G4.7 Detonators

German group 1 millisecond detonators were used, furnished with an extra insulating

hose and 6 m initiating cables. The electric resistance of each detonator was 1.8 ohm.

Two detonators of the same number were used in each hole, placed in the second

cartridge from the top and the bottom of the hole.

G4.8 Coupling

The two detonators of each hole were coupled in two separate series which then were

coupled in parallel. The firing cable had a cross section of 2 · 2.5 mm2, and was led

through the gate shaft to the gate house. The length of the firing cable was 530 m.

Scotchlok coupling clips, type UR, and a pincer of type E9Y were used. Insulation

and coupling to the firing cable was done by tape and a silicone compound. The

blaster was type CI 160 VA.

All the charging and coupling work was carried out with great caution, in order not to

damage the initiating cables. The dismantling of the scaffolding had to be done very

carefully. All the woodwork and other floating objects were removed before the water

filling started.

G4.9 Water Filling

The tunnel between the gate shaft and the plug was filled with water through the

shaft, using two pumps of 1000 m3/h.


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G4.10 Air Pocket

Due to leakage of air during the water filling, the pressure of the air pocket had to be

maintained by compressed air from two compressors. Maximum air pressure was 10

bar. An air hose of diameter 50 mm was led through the shaft to the air pocket.

G4.11 Technical Data

Calculations Measured Plug volum 270 sm3 No. of 45 mm drillholes 159 No. of 102 mm drillholes 8 Amount of explosives 1200 kg 1100 kg Spesific charging 4.5 kg/sm3 4.0 kr/sm3

No. of detonators 318 Water filling volume 20 000 m3 19 000 m3

Volume of air used 6150 Nm3 3200 Nm3

Max. air pressure at water filling 9.9 bar 7.5 bar Max. air pressure at blasting 14.3 bar 12.5 bar Elevations Water level Lake fossevatn 620.0 m 620.0 m Gate house floor 624.3 m 624.3 m Gate threshold 501.0 m 501.0 m Water in the shaft 600.0 m 572.8 m * Upsurge in the shaft 624.2 m * Absolute upsurge in the shaft 51.4 m * Pressure against the gate Max. allowed 18.1 bar Max. pressure at blasting 17.7 bar 17.3 bar • = The water level was reduced during filling, due to leakages from the air pocket through the plug, and

water leakages through the gate.

Table G.2 Technical data


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Time used for the piercing round:

Charging and coupling 17 hours

Dismantling the scaffolding 27 hours

Water and air filling 17 hours

Firing 0

Total 61 hours

G4.12 Results

The water had an upsurge in the gate shaft of 10 cm below the gate house floor. This

was 20 cm lower than calculated.

The pressure against the gate was measured at 17.3 bar, corresponding to a calculated

pressure of 17.7 bar. The gate was designed for a pressure of 18.1 bar.

The piercing was successfully carried out, but within small margins. All the necessary experts were present, which should be the case for all piercings, and changes from the assumed factors were

APPENDIX G2. The Troll Field Shore Approach

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G2. THE TROLL FIELD SHORE APPROACH

G2.0 Introduction During the autumn of 1979 the Norwegian company, A/S Norske Shell discovered the Troll Field, the largest offshore gas field in Europe. The field is located 80 kilometres north-west of Bergen. Unprocessed gas from the field is transported through pipelines to Kollsnes in Øygarden Community, where it is processed in order to meet export specifications and then transported to the European continent. The shore approach for the pipelines is a subsea tunnel system about 4 kilometres long consisting tunnels, shafts, concrete works, installation of steel support cones for rise bundles and the final blasts. The final blasts at -160 m were the deepest underwater piercing in the world. The largest risk related to the project applied to the final blasts. To solve this, a number of new solutions regarding drilling, blasting, testing and control systems were introduced. Hence, it took 2 years of engineering and production planning before the physical work started. This, together with the comprehensive safety steering system, the visible management of Norske Shell in safety matters, and the persistence of Norwegian crews combined with the flexibility in engineering, is the main reason for the success of the project.

Figure G.5 General shore approach system.


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G2.1 The Tunnel System The tunnel system includes one main tunnel, one secondary tunnel, branching tunnels in the piercing area, three shafts with technical installations and three final blasts and concrete plugs. The main tunnel has a decline 1:7 for 1800 metres and then goes horizontal for approximately 2 km. In the piercing area, the main tunnel branches into three parts. In each branch there is a compartment with a shaft. The shafts are 30 - 35 m long and have a 32 m2 cross section. From the shafts, the piercing would be performed 160 – 170 m below the sea level. It is not known that underwater piercing has been done at such depth previously.

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Figure G.6 General layout of the tunnel system. I = from the Troll Field, E=Export

to Europe. The final blast had a minimum cross section on 34 m2, and a circular and conical shape. The slope of the walls was 3.6° from the vertical. The seabed was provided horizontal with a layer of moraine.


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After the final blasts the prefabricated riser bundles were to be placed and concreted. The riser bundles were placed into steel support cones that were concreted in the shafts before the final blasts. It was a main requirement that the whole system was waterproof after the installation of the riser bundles.

G2.2 Design Parts of the final blasts had to be drilled before the steel cones were assembled. After the charging and coupling of the final blast, the scaffolding had to be disassembled. This means that a considerable quantity of work had to be performed in the shafts after charging the final blasts. The final blast was planned as follows: Length 5 – 7 m Bottom of drillhole to seabed 0.6 m Diameter of charged holes 51 mm Diameter of large holes 150 mm Explosives Extra Rubber Dynamite (Dyno Nobel) Detonators Non-el Drill plan Back-up for all holes Spacing of contour holes 30 cm Number of charged holes 236 Number of large holes 8 Minimum cross section 34.0 m2 Maximum cross section 43.0 m2


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The shaft and the shaft compartment were precompressed before the blasts. One problem with the design was the transport of blasted rock through the cones. To solve this, the duration of the blast was increased so that it became a sequential blast and mass transport. This also favoured the quality of the contour, but increased the risk to destroy charges because of detonation shocks or gas pressure from neighbouring holes. The following measuring system was used to control the blast and the environment: • = Vibration measurements at different levels in the shaft • = Measurements of pressure in the precompressed compartment and shafts • = Measurements of deformation in the concrete and steel assemblies.


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Figure G.7 Side view of a piercing plug.


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G2.3 Control The rock mass was pregrouted and the working conditions were better than average. The drilling was done with a specially designed hydraulic drilling rig. The charges for all holes were prefabricated at a level in the shaft during the charging. The drilling were perform with a maximum deviation of 1° and the holes had no observed deviation. The blasting was inspected with an ROV from the sea right after blasting. Except for the first blast, the contour was according to design without cracks or overbreak. Measurements of vibration and pressure progress in the precompressed compartment and shafts showed results according to plans. Insignificant deformation was measured. There were no surface damage to the cones or the steel support cones for the rise bundles. The two last blasts were changed regarding to the drilled length and the distance to the seabed was reduced to 40 cm. After assembling the rise bundles and pumping water out of the compartments, all three shafts were dry and waterproof. The conclusion from the Troll Field piercing is that strict control of the security leads to a high degree of safety, good precision and high productivity.

Underwater Tunel Piercing

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