NTNU Hard Rock Tunnel Boring Thesis

Geology and Site Investigations
Hard Rock Tunnel Boring

Doctoral theses at NTNU 1998:81
Amund Bruland
Vol. 5 of 10
NTNU Trondheim
Norwegian University of
Science and Technology
Doctoral thesis
for the degree of doktor ingeniør
Faculty of Engineering Science
and Technology
Department of Civil and
Transport Engineering
PREFACE 1
0 GENERAL 3
0.1 Project Reports about Hard Rock Tunnel Boring 3
1 ROCK MASS BOREABILITY 6
1.0 Introduction 6
1.1 Rock Drillability 7

1.11 Introduction 7
1.12 The Drilling Rate Index DRI 7
1.13 Rock Abrasiveness 8
1.14 The Cutter Life Index CLI 9
1.15 Rock Mineral Content 10
1.2 Rock Mass Fractures 12

1.20 Introduction 12
1.21 Classification of Fractures 13
1.22 Degree of Fracturing 14
1.23 Examples of Classification 17
1.24 Marked Single Joints 22
2 SITE INVESTIGATIONS 24
2.0 Introduction 24
2.1 Degree of Fracturing 27

2.11 Surface Mapping 27
2.12 Core Logging 32
2.13 The RQD Method 33
2.14 The Q-System 35
2.15 The RMR System 37
2.16 Seismics 39
2.17 Experience from Back-mapping 41
2.18 Rock Stress 42
2.2 Rock Drillability 45

2.21 Rock Sampling 45
2.22 Data from the Drillability Database 46
2.23 The Compressive Strength 48
2.24 The Point Load Index 50
2.25 Cerchar Abrasivity Index CAI 52
2.26 Vickers Hardness Number Rock 53
2.3 Rock Support 55
2.4 Engineering Geological Summary 56
2.5 Variability of the Rock Conditions 58

Continues next page
APPENDICES 61
A. Previous Editions 61
B. Research Partners 62
C. List of Parameters 63
D. The Q-System 66
E. The RMR System 72
F. Mean Orientation of Discontinuities 77

PREFACE
HARD ROCK TUNNEL BORING Geology and Site Investigations

Project Report 1D-98
The report is one of six reports about hard rock tunnel boring:
• 1A-98 HARD ROCK TUNNEL BORING Design and Construction

• 1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear
• 1C-98 HARD ROCK TUNNEL BORING Costs
• 1D-98 HARD ROCK TUNNEL BORING Geology and Site Investigations
• 1E-98 HARD ROCK TUNNEL BORING Performance Data and Back-mapping
• 1F-98 HARD ROCK TUNNEL BORING The Boring Process
In addition, HARD ROCK TUNNEL BORING Background and Discussion gives

general information about the basis of the above listed reports.
Combined with the other reports in the Project Report Series from the Department of
Building and Construction Engineering at NTNU, the reports present an updated and
systematised material 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 advance rate, cutter wear and excavation cost models also exist as a WINDOWS
programme.
The report is prepared by Amund Bruland and is part of his dr.ing thesis about
hard rock tunnel boring.
The reports listed above describes a comprehensive model developed at NTNU The
model covers the complete tunnel boring process from the early planning stage
1
PREFACE
through preinvestigations, time and cost estimates, tunnel excavation and finally ac-
quisition and treatment of experience data. The models and data presented in the
reports are meant to be a practical tool for owners, consultants and contractors,
more than a theoretical analysis of the tunnel boring process.
The project has been granted financial support by our external research partners, see
list in Appendix.
For reference, registration and similar, we ask for the following:
NTNU-Anleggsdrift (1998): Project Report 1D-98 HARD ROCK TUNNEL

BORING Geology and Site Investigations.
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 47 37 Fax +47 73 59 70 21

e-mail amund.bruland@bygg.ntnu.no
Internet http://www.bygg.ntnu.no/batek/batek.htm
2
0. GENERAL 0.1 Project Reports about Hard Rock Tunnel Boring
0.1 PROJECT REPORTS ABOUT HARD ROCK TUNNEL BORING
1D-98
The report provides methods and data to be used during preinvestigations and
back-mapping of engineering geological properties of importance to time con-
sumption and excavation costs for hard rock tunnel boring.
The report treats the following items:
• Chapter 1: Rock mass properties for the time and cost estimation models devel-
oped at the Department of Building and Construction Engineering at NTNU.
• Chapter 2: Describes and evaluates various methods of engineering geological
site investigations.
Project Report 1D-98 is partly based on the project reports 1-76, 1-79, 1-83, 1-88 and
1-94, all published by the Department of Building and Construction Engineering at
NTNU. The report presents updated and revised information from the previous re-
ports as well as experience from recent tunnelling projects and information published
by others.
Appendix A shows a list of previous editions of the HARD ROCK TUNNEL

BORING report.
Other Reports
The Project Report 1A-98 HARD ROCK TUNNEL BORING Design and Con-
struction describes general design parameters such as tunnel profile, tunnel inclina-
tion and curve radius. Some features of various tunnel types like water, sewage, road
and rail tunnels are treated. Transport, ventilation and other necessary service systems
are presented.
The Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and
Cutter Wear provides methods and necessary data for estimation of time consump-
3
tion and cutter wear for tunnel boring. Geological parameters and machine factors of
significance for the penetration rate and the cutter wear are presented briefly.
The Project Report 1C-98 HARD ROCK TUNNEL BORING Costs presents
models and data for estimation of tunnel excavation costs and total construction costs.
The Project Report 1E-98 HARD ROCK TUNNEL BORING Performance Data
and Back-mapping covers follow-up procedures and collecting of performance data
from tunnel boring projects. Engineering geological back-mapping is treated in detail.
Project Report 1F-98 HARD ROCK TUNNEL BORING The Boring Process
covers rock breaking and chipping, machine factors affecting performance, boring in
fractured rock mass, and various types of cutter wear.
Use of the Estimation Models
The estimation models are aimed at being used through several stages in a project:
• Preliminary and feasibility studies

• Project design and optimisation
• Tendering and contract
• Construction
• Possible claims.
The estimation models for Hard Rock Tunnel Boring should be used with care. Com-
bined with other estimation models in the Project Report Series from the Department
of Building and Construction Engineering, the Hard Rock Tunnel Boring reports pro-
vide a reliable and practical tool to be used for:
• Estimating net penetration rate and cutter wear

• Estimating time consumption and excavation costs, included risk
• Assess risk with regard to variation in rock mass boreability or machine
parameters
• Establish and manage price regulation in contracts
• Verify machine performance
• Verify variation in geological conditions.
4
Background
The estimation models are based on job site studies and statistics from tunnelling in
Norway and abroad, including more than 35 job sites and more than 250 km of tunnel.
The data have been systematised and normalised. The results are regarded as being
representative for well organised tunnelling.
A more detailed treatment of the background and the basis for the Hard Rock Tunnel
Boring estimation models is found in HARD ROCK TUNNEL BORING Back-
ground and Discussion.
5
1. ROCK MASS BOREABILITY 1.0 Introduction
1.0 INTRODUCTION
With regard to boreability, the rock mass consists of intact rock and planes of weak-
ness.
In the estimation models for penetration rate and excavation costs presented in the
Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and
Cutter Wear and the Project Report 1C-98 HARD ROCK TUNNEL BORING
Costs, the rock mass properties concerning tunnel boring are described by:
• The rock mass type of fracturing

• The rock mass degree of fracturing
• The orientation of the fracturing system(s)
• The drillability of the intact rock
• The abrasiveness of the intact rock
• The porosity of the intact rock.
The influence of the rock stress should also be included in the estimation models.
Field data indicates that in favourable stress situations, the net penetration rate will
increase considerably. See also Section 2.16.
The characterisation of the rock mass boreability must always bear in mind that dur-
ing the site investigations one is not able to map or investigate the complete rock mass
to be excavated. Hence, one must rely on interpretation and extrapolation of informa-
tion from selected and available spots along the tunnel route.
6
1. ROCK MASS BOREABILITY 1.1 Rock Drillability
1.1 ROCK DRILLABILITY
1.11 Introduction
The intact rock drillability is evaluated by the Drilling Rate Index DRI and the Cutter
Life Index CLI. The test methods are described in the Project Report 13A-98
DRILLABILITY Test Methods.
The compressive strength and the point load strength are useful supplementary pa-
rameters, but are found to be less significant to determine the rock drillability than the
Drilling Rate Index DRI. Hence, these parameters are not used as direct input pa-
rameters in the estimation models. See also Chapter 2.
The porosity should be measured for rock types with porosity higher than approxi-
mately 2 %. The influence of the porosity on the DRI has been found to be negligible.
1.12 The Drilling Rate Index DRI
The DRI has been selected as the drillability parameter of the intact rock since it is
believed to give a good representation of the rock breaking process under a cutter.
The cutter edge indentation into the rock surface is believed to be very important to
achieve efficient transfer of the forces and chipping. Hence, the drillability parameter
should include the rock resistance to indentation or the rock surface hardness.
Furthermore, the drillability parameter should simulate the dynamic rock breaking
and chip forming process which is believed to consist of tensile crack formation as the
main part.
The Drilling Rate Index DRI of a rock is the brittleness value adjusted for surface
hardness, and is an indirect measure of the required breaking work, where:
• The Brittleness Value S20 expresses the amount of energy required to initiate
cracks and crush the rock. The test is believed to crush the rock mostly by induced
tension.
7
• The Sievers' J-value SJ expresses the depth the cutter can be thrust into the rock,
in other words how efficiently the rock brittleness (i.e. the cracking tendency) can
be utilised.
1.13 Rock Abrasiveness
1
Abrasion Value Steel, AVS
50
4
2
40
5
30
20
1 Q = 0 - 10%
2 Q = 11 - 20%
10 3 Q = 21 - 40%
4 Q = 41 - 70%
5 Q = 71 - 100%
0
0 20 40 60 80
Abrasion Value, AV
Figure 1.1 Relation between the Abrasion Values AV and AVS for varying quartz
content Q in percent. Data from the Project Report 13C-98
DRILLABILITY Statistics of Drillability Test Results.
Line Equation r2
1 y = 1.4743x1.0387 0.70
0.8611
2 y = 1.4836x 0.79
3 y = 1.5491x0.8183 0.75
0.7796
4 y = 1.4588x 0.59
0.5604
5 y = 3.1383x 0.27
8
The Abrasion Value measures the time dependent abrasiveness of the rock, and is
determined for two types of cutter material:
• The Abrasion Value AV using tungsten carbide as test piece.

• The Abrasion Value Steel AVS using a cutter steel as test piece.
Hence, the rock abrasiveness on cutters with carbide inserts is expressed by the AV,
while the rock abrasiveness on cutter rings of steel is expressed by the AVS.
The AV and AVS are measured using crushed rock powder less than 1 mm and is
meant to simulate the abrasion on the cutter rings from the crushed rock powder in the
kerfs and from the chips and broken rock flowing past the cutters from the rock face.
The relation between the AV and the AVS is shown in Figure 1.1. The curves are
based on approximately 800 parallel samples. The correlation coefficient shows that
the relations should be used with caution, especially for rock types with quartz con-
tent of more than 40 %.
1.14 The Cutter Life Index CLI
The Cutter Life Index CLI expresses the cutter ring life in hours and is calculated on
the basis of the Abrasion Value Steel AVS and Sievers' J-value SJ by means of the
empirical relation in [1.1].
0.3847
 SJ 
CLI = 13.84 ⋅   [1.1]
 AVS 
The CLI is designed to comprise what is believed to be the two most important wear
factors for cutter rings of steel.
The SJ represents the surface hardness of the rock, as well as the wear mode and the
wear pattern of the cutter ring.
9
A low SJ will mainly result in point abrasion on the cutter ring edge due to low in-
dentation while a high SJ will result in side abrasion as well as point abrasion of the
cutter ring. It has been observed that a low SJ results in less crushed rock powder in
the kerf than a high SJ does. It is believed that the less amount of rock powder in the
kerf results in a somewhat different and more abrasive wear mode than more crushed
rock powder in the kerf does, since the cutter ring will have more direct contact with
the intact rock surface.
1.15 Rock Mineral Content
When normalising the field data to make the estimation model in the Project Report
1B-98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear, it was
necessary to include a correction factor for the rock content of hard and abrasive min-
erals, represented by the quartz content. The quartz content should also include min-
erals like epidote and garnet.
1 .6 1 M ic a s c h is t
2
k Q M ic a g n e is s
1 .4 G n e is s
G r a n itic G n e is s
1 .2 G r a n ite
2 O th e r ro c k ty p e s
1 .0
0 .8
1
0 .6
0 .4
0 2 0 4 0 6 0 8 0 1 0 0
Q u a rtz c o n te n t, %
Figure 1.2 Correction factor for rock quartz content. From the Project Report 1B-
98 HARD ROCK TUNNEL BORING Advance Rate and Cutter Wear.
10
The possible explanation of the need of a correction factor for quartz content is an
assumed difference between the abrasion under a cutter and the abrasion test in the
laboratory. Since the pressure under the cutter is larger than under the test piece in the
laboratory, a different sieve curve of the rock powder is supposed. Under the cutter,
the softer minerals are crushed to such an extent that they do not function as mineral
grains any longer. The hard minerals become overexposed, which again results in a
higher ring abrasion. This explanation is applicable for the rock types of Group 2 and
for the rock types of Group 1 with a quartz content of more than 30 %.
The progress of the correction factor for the rock types of Group 1 with quartz content
of less than 30 % has not been sufficiently explained.
11
1. ROCK MASS BOREABILITY 1.2 Rock Mass Fractures
1.2 ROCK MASS FRACTURES
1.20 Introduction
Planes of weakness or discontinuities in the rock mass contribute considerably to the

net penetration rate and cutter wear for tunnel boring. In this context, "planes of
weakness" means planes with little or no strength along the planes. The influence of
the planes of weakness on the penetration rate depends on several factors. Table 1.1
shows the most important factors.
Rock Mass Properties Machine Factors
• Type, frequency and orientation of the • Thrust level

planes of weakness
• Cutter size and type
• Rock drillability
• Average cutter spacing
• Machine diameter
• Torque capacity
Table 1.1 Important factors deciding the influence of the planes of weakness on the
penetration rate and the cutter wear.
The interaction between the rock mass properties and the machine factors is described
in the Project Report 1F-98 HARD ROCK TUNNEL BORING The Boring Pro-
cess.
The use of the term "Fracture" instead of "Discontinuity" intends to exclude two of
the discontinuity types in the ISRM1,2 recommended definition: The weakness zones
and faults. The ISRM definition is as follows: "The general term for any mechanical
discontinuity in a rock mass having zero or low tensile strength. It is the collective
term for most types of joints, weak bedding planes, weak schistosity planes, weakness
zones and faults."
1
E. T. Brown (ed.): Rock Characterization Testing and Monitoring - ISRM Suggested Methods, Pergamon
Press 1981.
2
ISRM - International Society for Rock Mechanics.
12
1.21 Classification of Fractures
The estimation models in the Project Report 1B-98 HARD ROCK TUNNEL
BORING Advance Rate and Cutter Wear group the planes of weakness as follows.
• Systematically fractured rock mass.

Parallel oriented joints.
Parallel oriented fissures.
Foliation planes or bedding planes.
• Marked Single Joints.
• In addition, Weakness zones and Faults are registered, but not included in the es-
timation model for net penetration rate.
The classification of weakness planes includes:
• Type of planes with regard to persistence, openness or aperture, clay or mineral

filling, roughness and planeness.
• Joint or fissure spacing (average distance between the planes of weakness).
• Orientation to the tunnel axis.
The basic features of the various types of weakness planes included in the model are
given below.
• Joints (denoted Sp): Includes continuous joints that can be followed all around
the tunnel profile. They can be open (e.g. bedding in granite) or filled with clay or
weak minerals, e.g. calcite, chlorite or similar minerals.
• Fissures (denoted St): Includes non-continuous joints or fissures (can only be

followed partly around the tunnel profile), filled joints with low strength, and fo-
liation and bedding plane fissures, e.g. as in mica schist or mica gneiss.
• Homogenous Rock Mass (denoted Class 0): Includes massive rock without
joints or fissures (may occur in intrusive dikes, sills, batholiths, etc.). Rock mass
with filled joints of high strength (e.g. joints healed with quartz, epidote, etc.) may
approach Class 0.
13
• Marked Single Joints (denoted ESP or MSJ): Includes marked discontinuities

in the rock mass such as large exfoliation joints and feather joints. They may be
completely open, lead water or be filled with clay. Minor faults filled with gouge
may be classified as particularly prominent single joints.
The grouping of fissures and joints is not exact and objective, but somewhat depend-
ant on the person doing the mapping. Since the classification system was developed,
experience from back-mapping of bored tunnels shows that the major part of the rock
mass fracturing is classified as Fissures or Marked Single Joints. Systematic joints
occur for shorter sections only.
1.22 Degree of Fracturing
The degree of fracturing in systematically fractured rock mass is grouped into classes
for practical use when mapping (see Table 1.2). The classes include both the spacing
between and the type of weakness planes.
Fracture Class Spacing between the Grouping of Classes

(Joints = Sp / Fissures = St) Planes of Weakness with Regard to Spacing
af [cm] [cm]
0 ∞ 240 - ∞
0-I 160 120 - 240
I- 80 60 -120
I 40 30 - 60
II 20 15 - 30
III 10 7.5 - 15
IV 5 4 - 7.5
Table 1.2 Fracture classes with spacing between the planes of weakness.
The fracturing frequency or the spacing may be given as Class or directly in cm. Us-
ing Classes is recommended since giving the spacing in cm may be misleading with
regard to accuracy. The Classes listed in Table 1.2 is not sufficient to describe the
14
fracture spacing. For practical mapping one adds a "+" or a "-" to the Class and also
uses the combination of adjacent Classes.
Table 1.3 shows the notation used for a more detailed classification in the Class II and
Class I range.
Average spacing between the Fissures

Fissure Class
af [cm]
40 St I
35 St I+
30 St I - II
25 St II-
20 St II
Table 1.3 Subdividing of fracture classes with spacing between the planes of weak-
ness.
Nature is, however, far more varied than such a simplified classification can express.
It is only in schist or similar rock types that the spacing of fractures may be found to
have a constant value. The spacing between the planes of weakness is measured per-
pendicularly to the planes and given as an average for some distance.
In weak, schistose rock, e.g. mica schist, it is difficult to distinguish between the
schistosity of the rock and fissures along the schistosity planes, see Figure 1.3. It is
then essential to not include the effect of the schistosity twice, i.e. both in the drilla-
bility parameter and as an addition for planes of weakness. This is particularly im-
portant when compressive strength is used as a parameter for drillability.
In schistose rocks with high tensile strength parallel to the schistosity planes, e.g.
mica gneiss, cross joints or fissures will often give the highest penetration rate addi-
tion. Surface mapping of cross joints or fissures in this type of rock mass is difficult.
Core drilling may then be used to find the degree of fracturing. The orientation to the
tunnel axis will still be difficult to determine.
15
F is s u r e s a lo n g th e s c h is to s ity ?
F is s u re s
a lo n g th e
s c h is to s ity
p la n e s
R o c k s a m p le
C r o s s jo in ts o r fis s u r e s
Figure 1.3 Schistosity fissures and cross jointing in a schistose rock - e.g. mica
schist, quartz schist, etc.
Folding may also complicate the picture. In such cases, one has to make an approxi-
mate evaluation of the degree of fracturing.
When more than one set of weakness planes is present, as shown in Figure 1.4, each
set is mapped individually and the total fracturing factor is estimated as follows:
 n 
k s −tot =  ∑ k si  − (n − 1) ⋅ 0.36 [1.2]
 i =1 
ks-tot = total fracturing factor

ksi = fracturing factor for set no. i
n = number of fracturing sets.
The fracturing factor ks is found in the Project Report 1B-98 HARD ROCK
TUNNEL BORING Advance Rate and Cutter Wear.
16
F ra c tu re
s e t 1
F ra c tu re
s e t 2
a f1
f2
a
Figure 1.4 Rock mass with two sets of systematic fractures.
When the rock mass fracturing is difficult to separate into sets, one may use an aver-
age spacing and an average angle for the total fracturing. In most cases this will be
sufficiently accurate. The accumulated volumetric area of fractures in a given rock
volume may also be used to estimate the degree of fracturing. However, when the
orientation of fractures seems to be random, it is recommended to just count the frac-
tures in a given rock volume and estimate the degree of fracturing on that basis. The
largest source of error will be the averaging of the angle between the tunnel axis and
the planes of weakness.
1.23 Examples of Classification
The Figures 1.5 - 1.13 show examples of evaluation of the degree of fracturing. The
classification examples are first and foremost relevant when one set of weakness
17
planes predominates. If more than one set occur, the influence from each set can be
estimated, but for the combined influence, sound judgement must be exercised.
Figure 1.5 Mica gneiss, Fissure Class I-.
Figure 1.6 Granitic gneiss, Fissure Class 0-I. The fissures are exposed to iron oxide
corrosion.
18
Figure 1.7 Granitic gneiss, Joint Class 0-I.
Figure 1.8 Drill cores of phyllite with Fissure Class IV+. Such a high fracturing
class is not common except in shear zones or similar.
19
Figure 1.9 Phyllite with Fissure Class III. The planes are uneven and rough.
Figure 1.10 Mica schist with Fissure Class IV.
20
Figure 1.11 Mica gneiss with Fissure Class III+.
Figure 1.12 Mica schist with Fissure Class II.
21
Figure 1.13 Gneiss with Fissure Class II-.
1.24 Marked Single Joints
Single joints are recorded individually. They are marked discontinuities in the rock
mass. As the name indicates, they will appear as a singular phenomenon at the tunnel
level, even if they belong to a systematically occurring joint set on a larger scale.
Marked Single Joints may result in very high net penetration rates locally. Due to the
risk of stability problems and damage to the cutters and the cutterhead, one should
evaluate the effect carefully and consider possible delays if the Marked Single Joints
may result in mixed face conditions, heavy rock support, water inflows, etc.
22
Figure 1.14 Marked Single Joints (mica layers) in granitic gneiss.
Figure 1.15 Marked Single Joint in the tunnel (mica layer from Figure 1.14).
23
2. SITE INVESTIGATION 2.0 Introduction
2.0 INTRODUCTION
The rock mass boreability is far more essential to performance and costs for tunnel
boring than for drill and blast tunnelling. The geological and engineering geological
site investigations must therefore be adjusted to the tunnelling method.
The estimation models in the Project Report 1B-98 HARD ROCK TUNNEL
BORING Advance Rate and Cutter Wear emphasise a numerical description of the
geological conditions so that the rock mass boreability may be quantified as good as
possible by the site investigations.
The site investigations should provide a basis for estimating:
• The net penetration rate in m/h.

• The cutter life in h/cutter.
• The gross advance rate in m/week.
• The total construction time.
• The total construction costs.
• The risk with regard to time consumption and excavation costs due to variation in
the expected geology.
To be able to use the estimation models mentioned above, the site investigations must
as a minimum provide the following:
• Rock mass fracturing, preferably described by Fracturing Class and orientation.

Data from other characterisation systems may be interpreted and adapted to the
estimation models.
• Drillability of the intact rock, preferably given as DRI, CLI and quartz content.
Other parameters like compressive or tensile strength, CAI or VHNR may be used
as a basis for estimation of the input parameters of the estimation models.
According to the estimation models, the gross advance rate is estimated without in-
cluding the rock support that necessitates stop in the boring process. The excavation
cost model do not include any rock support. The extra time consumption and costs
related to rock support must be estimated separately. Hence, the site investigations
must include:
24
• Fault zones and zones with mechanically weak rock.

• The rock mass stress situation.
• The rock mass water conditions.
• Other features involving extra time and costs.
Experience from the back-mapping of bored tunnels show that the geological condi-
tions with regard to the rock drillability and the rock mass degree of fracturing will
have a quite large variation along the tunnel axis, even for short tunnel sections, see
Figure 2.1.
160
Cumulated length of Fracture Classes, m
140
120
100
80
60
40
20
0
St 0+ St 0-I St I- St I St I+ St I -II
Fracture Class
Figure 2.1 Variability of the rock mass degree of fracturing for a 410 m long tunnel
section.
Since the rock mass shows such variation, one consequence is that it is not a prerequi-
site that the classification of the rock mass fracturing must use drill cores reaching the
tunnel level or that all samples for laboratory testing are collected at the tunnel level.
The rock mass boreability should describe average values for larger parts of the tunnel
to be bored.
25
The mapping of the degree of fracturing and the rock sampling will in most cases be
in the form of available or selected locations or points along the tunnel alignment.
When the field mapping data are to be aggregated and the tunnel divided into geo-
logical sections with corresponding rock mass parameters, statistics must be used with
caution, leaving the final assessment of the parameters to the judgement of the per-
son(s) who performed the mapping.
A general problem in most site investigations for tunnel projects is the extrapolation
of surface mapping to the tunnel level. There are several approaches to be used in the
extrapolation, but basically on must accept an element of uncertainty in the predicted
geological conditions at the tunnel level.
The main purpose of the site investigations must be to establish a sufficiently accurate
model of the geological conditions to be able to estimate the excavation costs and the
time consumption within acceptable levels of uncertainty.
The extent of the field mapping must be considered for each project. Where the cost
difference between tunnel boring and drill and blast is expected to be low, or where
the rock mass is especially difficult to bore, an extensive mapping should be carried
out.
The site investigations should not focus on the engineering geological parameters
only. It is strongly recommended that the general geology and the geological history
of the site area are incorporated in the site investigations. This will help to interpret
and extrapolate the mapping data.
26
2. SITE INVESTIGATION 2.1 Degree of Fracturing
2.1 DEGREE OF FRACTURING
2.10 Introduction
The rock mass fracturing is by far the most important factor in the TBM performance
estimation. Hence, the site investigations should focus more on the overall rock mass
fracturing than on the drillability of the intact rock.
This means that the site investigations for TBM projects should put more emphasis on
the rock mass fracturing outside fault zones, crushed zones and similar than what has
been the standard procedure until now. The properties of the intact rock must of
course be tested in the laboratory as a part of the site investigations.
It is recommended to use the Fracture Classes and the orientation of the planes of
weakness when carrying out the site investigations with regard to estimation of the
penetration rate for a TBM project. However, the RQD, Q and RMR values are fre-
quently used to describe the rock mass. The tentative relations between the three pa-
rameters and the Fracture Class are discussed briefly below.
2.11 Surface Mapping
Surface mapping is an easy and relatively cheap way to map the degree of fracturing
and the orientation of the planes of weakness. However, it may require some experi-
ence to extrapolate the results of the surface mapping to the tunnel level. It is there-
fore strongly recommended that the person doing the surface mapping should have
experience from back-mapping of bored tunnels.
The surface mapping should cover a corridor along the tunnel route. The width of the
corridor should be proportional the depth of the tunnel. The practical mapping is
performed as mapping at points or concentrated areas in accordance with the selected
corridor. When the data from the mapping later are aggregated and averaged, it is im-
portant to consider the varying distance between the mapping locations and not give
each location the same weight if they represent various distances along the tunnel.
The quality of observations made at each location must be evaluated when data are
aggregated and averaged.
27
The following procedures may be applied at each mapping location:
• Determine the co-ordinates of the location according to the map.

• Describe the location with regard to the quality and reliability of the mapping
results, e.g. flat outcrop, fresh road cut, etc.
• Determine the rock type and collect a hand specimen.
• Evaluate the location with regard to the possible collection of rock sample(s) for
laboratory testing.
• Determine the structure of the rock, e.g. foliation, banding, etc.
• Group the discontinuities of the rock mass into sets of Fissures or Joints,
Marked Single Joints, Faults, etc.
• Determine the average spacing for the Fissure and Joint sets and the corre-
sponding Fracture Class.
• Evaluate the frequency of possible Marked Single Joints.
• Describe possible faults with regard to type, size, filling material, etc.
• Measure the orientation of as many discontinuities as possible.
• Take at least one photograph at the mapping location.
It is an advantage to record the degree of fracturing at fresh rock surfaces, e.g. nearby
tunnels, road cuts, river cuts, cliffs and similar.
The degree of weathering plays an important role in the classification of fractures ob-
served at the surface. Even at a low degree of weathering, fissures and joints will be-
come overexposed. A general rule is that what seems to be joints at the surface will
occur as fissures at the tunnel level.
In a hilly or rolling terrain it is important to remember that the hills usually represent
stronger and less fractured rock mass than the valleys or troughs.
Some joints and fissures are surface phenomena, e.g. exfoliation joints in granite and
fissures from frost shattering in schistose rock types. These planes of weakness do not
extend to the depth, and should not be included in the estimates.
When mapping in road cuts and drill and blast tunnels one will get an overexposed
impression of the degree of fracturing. Fractures categorised as fissures in a bored
28
tunnel and at the surface will often appear as joints in a drill and blast tunnel or in a
cut. In addition, the fracturing frequency may be overestimated due to blast damage.
Aerial photos are a useful aid to get an impression of the general structure of the rock
mass, Marked Single Joints or larger fractures and crushed zones.
Seismic investigations to determine the degree of fracturing of the surface rock have
been used. Seismic methods are, however, not sufficiently developed to determine the
degree of fracturing at depth.
Orientation
D ip d ir e c tio n to th e r ig h t h a n d D ip d ir e c tio n to th e le ft h a n d
s id e o f th e s tr ik e d ir e c tio n s id e o f th e s tr ik e d ir e c tio n
N N
S tr ik e d ir e c tio n
a s= 2 0 g
a d 1 0 0 g
D ip a d D ip
d ir e c tio n a s= 2 2 0 g d ir e c tio n
g
1 0 0
D ip D ip
p o le a f= 3 5 g p o le a f= 3 5 g
S tr ik e
d ir e c tio n
L o w e r h e m is p h e r e L o w e r h e m is p h e r e
N o ta tio n : N o ta tio n :
s tr ik e /d ip = a s /a f = 0 2 0 g /3 5 g s tr ik e /d ip = a s /a f = 2 2 0 g /3 5 g
d ip d ir e c tio n /d ip = a s + 1 0 0 g /a f = 1 2 0 g /3 5 g d ip d ir e c tio n /d ip = a s - 1 0 0 g /a f = 1 2 0 g /3 5 g
Figure 2.2 Options for strike and dip measurements and notation.
The orientation of the planes of weakness may be found by strike and dip measure-
ments or measurements of dip direction and dip. It is recommended that the strike, the
29
dip direction and the dip be measured in the full 360° or 400g system, not in a 180° or
200g system. The strike direction must be decided according to a chosen practise with
two options: The dip direction is always either to the right or the left-hand side of the
strike direction, see Figure 2.2.
The strike and dip measurements are preferably presented as plots in polar co-
ordinates of the discontinuity normal poles of the lower hemisphere, see the example
in Figure 2.3. To estimate the "mean" orientation of each set, visual judgement is suf-
ficiently accurate. When a set of discontinuities consists of several observations of
strike and dip, it may be advantageous to calculate the resultant normal by using the
direction cosines, see Appendix F.
The normal to a discontinuity is calculated according to [2.1] - [2.4], using polar co-
ordinates. When the pole of the "mean" normal for each set of discontinuities has been
estimated, the resulting strike and dip of each set is calculated by [2.5] - [2.8].
α sn = α s + 300 g ( g ) (dip direction to the right-hand side of the strike direction) [2.1]
α sn = α s + 100 g ( g ) (dip direction to the left-hand side of the strike direction) [2.2]
α dn = α d + 200 g ( g ) (dip direction and dip angle measurements) [2.3]
α fn = 100 g − α f (g ) [2.4]
αs = measured strike direction (g)

αd = measured dip direction (g)
αf = measured dip angle (g)
αsn = strike direction of the normal to the discontinuity plane (g)
αdn = dip direction of the normal to the discontinuity plane (g)
αfn = dip angle of the normal to the discontinuity plane (g)
30
g
0
S e t 3
S e t 1
g g
3 0 0 1 0 0
S e t 2
S e t 3
g
2 0 0
= e s tim a te d p o le
Figure 2.3 Plot of the normals to discontinuity planes using the lower hemisphere.
The "mean" normal is estimated by visual judgement.
α s = α sn − 300 g ( g ) (dip direction to the right-hand side of the strike direction) [2.5]
α s = α sn − 100 g ( g ) (dip direction to the left-hand side of the strike direction) [2.6]
31
α d = α dn − 200 g ( g ) (dip direction and dip angle measurements) [2.7]
α f = 100 g − α fn (g ) [2.8]
When the strike or dip direction and the dip of each set of discontinuities are known,
the orientation to the tunnel axis may be calculated according to [2.9] or [2.10].
α = arcsin ( sin α f ⋅ sin ( α t - α s )) (g ) [2.9]
α = arcsin ( sin α f ⋅ sin ( α t - α d + 100 g )) (g ) [2.10]
αt = direction of the tunnel axis.
2.12 Core Logging
In terrain with extensive overburden of loose deposits or heavily surface weathered

rock, drill cores may be used to determine the type of weakness planes, and the degree
of fracturing at the tunnel level may be estimated. However, it is difficult to determine
the orientation of the weakness planes from cores. Furthermore, planes of weakness,
which do not intersect the borehole, will not be recorded.
The best way to interpret the rock mass degree of fracturing from cores is to use aver-
age values for a longer section of the borehole, e.g. 10 m.
• Group the fractures into sets if possible.

• Count the number of fractures in each set for the section length.
• Estimate the angle of intersection between the fracture set and the core.
• Find the average spacing for each set by [2.11].
32
l c ⋅ sin α c
af = ( m) [2.11]
nf
af = average spacing between fractures (m)

lc = length of core section (m)
αc = angle of intersection between the fracture set and the core axis (° or g)
nf = number of fractures of the given set in the core section.
The angle between the planes of weakness and the tunnel axis must be found by other
means, e.g. general geology of the area, surface mapping or video logging of the
corehole.
If the rock sampling for laboratory testing may be done without the use of core drill-
ing, the investigation holes may be drilled by cheaper and faster methods like hammer
drilling. The degree of fracturing and the orientation of the fractures may be found by
video logging of the borehole.
An investigation hole does not necessarily need to reach the tunnel level, but it must
at least reach into fresh rock mass to avoid weathering phenomena.
2.13 The RQD Method
D. U. Deere introduced the RQD method (Rock Quality Designation) in 1963. The
RQD is usually the percentage recovery of core in lengths greater than 100 mm
(originally it was lengths greater than twice the core diameter).
RQD =
∑ Length of core pieces > 100 mm in length (%) [2.12]
Total length of core section
A first impression is that there should be a relation between the RQD and the Frac-
turing Class. But the use of RQD to determine the degree of fracturing may be mis-
leading. The method does not consider the average spacing of discontinuities, the ori-
enta-
33
ation of the weakness planes to the borehole or how many sets of fractures that are
present.
The Figure 2.4 shows examples of classification of RQD and Fissure Class. As indi-
cated by the figure, RQD values of 50 or less will most likely correspond to a Fracture
Class of II or higher. Transforming the RQD value into Fracturing Class will usually
give a conservative estimate of the degree of fracturing.
R Q D F is s u r e C la s s
> 1 0 0 m m > 1 0 0 m m > 1 0 0 m m > 1 0 0 m m > 1 0 0 m m
1 0 0 » S t II
< 1 0 0 m m < 1 0 0 m m
» 8 0 » S t II
< 1 0 0 m m
» 9 0 » S t I
< 1 0 0 m m
» 0 » S t III-IV
1 m
Figure 2.4 Classification of RQD and Fissure Class from core logging.
S. D. Priest and J. A. Hudson presented a correlation between RQD and discontinuity

spacing in 19761. The correlation is shown in Figure 2.5 and should be used with
caution.
1
S. D. Priest and J. A. Hudson: Discontinuity Spacings in Rock, Int. J. Rock Mech. Min. Sci. Vol. 13 1976,
pp 135 - 148.
34
R Q D
%
8 0
R Q D m a x
6 0
A v e r a g e c o r r e la tio n lin e
4 0
2 0
R Q D m in
0
0 2 0 6 0 2 0 0 6 0 0
M e a n d is c o n tin u ity s p a c in g , m m
Figure 2.5 Correlation of mean discontinuity spacing and RQD.
2.14 The Q-System
N. Barton, R. Lien and J. Lunde at the Norwegian Geotechnical Institute originally

presented the Q-system in 1974. The method has later been revised (1993). Basically,
the Q-System is designed for estimation of type and quantity of rock support.
The Q value of a rock mass is found by combining six parameters.
RQD J r J w
Q= ⋅ ⋅ [2.13]
J n J a SRF
35
RQD = Rock Quality Designation

Jn = relates to the number of joint sets
Jr = relates to the roughness of the most important joints
Ja = relates to the joint wall rock alteration and/or the joint filling material
Jw = relates to the water characteristics of the rock mass
SRF = relates to the stress condition in the rock mass.
The parameter values of the Q-system are given in Appendix D.
The Q value is not influenced by the orientation of the planes of weakness.
The possible weakness of the RQD as a parameter for the degree of fracturing is men-
tioned above. Of the other parameters included in the Q value, the Jn and Jr are
probably of most significance when comparing the Q-system to Fracturing Classes.
The Q value may vary from >400 as "Exceptionally good" to <0.01 as "Exceptionally
poor". Generally, a high Q value will correspond to a low degree of fracturing and
vice versa. The Q value may vary by more than 10 times for the same Fracture Class.
To give a more precise transformation of the Q value to the degree of fracturing, the
values of the background parameters must be known. Table 2.1 shows some examples
of Q value and Fracture Class, and that using the Q value directly to estimate the de-
gree of fracturing may be ambiguous.
Average Spacing RQD Jn Jr Ja Jw SRF Q Fracture Class

af (m)
0.5 100 4 4 1 1 1 100 St I(-)

0.25 100 2 4 1 1 1 200 St II-
0.4 80 4 3 1 0.66 1 40 St I
0.1 50 4 1 1 1 1 12.5 St III
0.1 30 9 0.5 2 0.66 1 0.55 St III
Table 2.1 Examples of classification by Q value and Fracture Class.
36
Figure 2.6 shows a rough comparison of Q values and Fissure Class. The comparison
is based on simulation of the Q value in addition to some few field data. When esti-
mating the Fissure Class from the Q value, caution must be exercised.
IV R Q
D
F is s u r e C la s s
i.e a n
. J d
r » J
J n
d e
a »
J c id
III
w
» S e s th
R F e
Q
» 1 val
u e
T h ,
m a e Q
jo r v a l
II ity u
o f e in f
th e lu
s ix e n c e
in p d b
u t y
p a th e
ra m
I e te
rs
0
0 .1 1 1 0 1 0 0
Q V a lu e
Figure 2.6 Rough comparison of Q value and Fissure Class.
2.15 The RMR System
Z. T. Bieniawski presented the Rock Mass Rating System or the Geomechanics Clas-
sification in 1973. The RMR System is designed to provide a general rock mass de-
scription to be used for various purposes, e.g. rock support or rock mass strength.
The RMRbasic value varies from 0 to 100 and is found by summing the ratings of five
rock mass parameters. It is also recommended to use an optional 6th parameter to ad-
just for joint orientation to a given construction or excavation, resulting in RMRadj.
37
RMR = RMR1 + RMR2 + RMR3 + RMR4 + RMR5 (+ RMR6 ) [2.14]
RMR1 = The strength of the intact rock (compressive or point load strength).
RMR2 = RQD Rock Quality Designation.
RMR3 = The average spacing of discontinuities.
RMR4 = The condition of discontinuities.
RMR5 = The groundwater conditions.
RMR6 = The orientation of discontinuities relative to the construction.
The parameter values of the RMR system are given in Appendix E.
The RMR value includes one important parameter for the degree of fracturing, the
average joint spacing. It also includes the rock strength, which may disturb the rela-
tion between the RMR value and the Fracture Class. Generally, a high RMR value will
correspond to a low degree of fracturing and vice versa.
The RMR value may vary quite much for the same Fracture Class. To give a more
precise translation of the RMR value to the degree of fracturing, the values of the
background parameters and field measurements must be known. In that case it is pos-
sible to estimate the degree of fracturing and the orientation of fractures with good
accuracy. Table 2.2 shows some examples of RMR value and Fracture Class, and that
using the RMR value directly to estimate the degree of fracturing may be ambiguous.
Average Rock RQD Joint Joint Ground- RMR Fracture

Spacing Strength Spacing Condition water Class
(m)
0.5 15 20 20 25 10 90 St I(-)
0.25 12 20 19 25 10 86 St II-
0.4 15 16 19 20 4 74 St I
0.1 7 10 14 12 10 53 St III
0.1 7 6 14 6 4 37 St III
Table 2.2 Examples of classification by RMR value and Fracture Class. The rock
mass is the same as in Table 2.1.
38
Figure 2.7 shows a rough comparison of unadjusted or basic RMR values and Fissure
Class. The comparison is based on simulation of the RMR value in addition to some
few field data. When estimating the Fissure Class from the RMR value, caution must
be exercised.
T h Q D
IV
R
e
F is s u r e C la s s
R M n d
a
R th e
va
lu p a
e
s
m
III
o s g o
c in
tly f d
d e is c
c id o n
T h
e d tin
in f e R M
b y u
lu e
II n c R v a
th
e
e d
itie
l
b y u e m
s
a ll o
in p r e o
u t r le
p a s
ra m s
I e te
rs
0
0 2 0 4 0 6 0 8 0
U n a d ju s te d R M R V a lu e
Figure 2.7 Rough comparison of RMR value and Fissure Class.
2.16 Seismics
Seismic velocity may be used as an indicator of the rock mass degree of fracturing.
Figure 2.8 shows common seismic velocities in Scandinavian soil and rock material.
The variability may be larger than indicated in the figure, and it must be emphasised
that seismic velocities must be used with great caution when the degree of fracturing
is to be estimated.
39
The variation in seismic velocity for a rock type in Figure 2.8 may represent variation
in the degree of fracturing, as low velocity indicates a high degree of fracturing and
high velocity indicates a sound rock mass with little fracturing. However, it is not
advisable to directly relate the variation in seismic velocity shown in Figure 2.8 to the
variation in degree of fracturing shown in Figure 2.10.
COMMON SEISMIC VELOCITIES IN SCANDINAVIA [m/s]

0 1000 2000 3000 4000 5000 6000 7000
WATER Water
Ice
TYPE OF SOIL Humus
Above ground- Peat
water level Clay
Sand
Gravel
Moraine
Clay
Below ground- Sand
water level Gravel
Moraine, loose
Moraine, hard
TYPE OF ROCK Sandstone
Limestone
Granite
Gneiss
Gabbro
Diabase
Figure 2.8 Seismic velocities in Scandinavia. From Project Report 19-88 UN-
DERWATER TUNNEL PIERCING.
The seismic velocities used to estimate the degree of fracturing must be measured for
relatively short sections (a few hundred metres or less) of otherwise uniform geologi-
cal conditions, especially with regard to rock type and absence of weakness zones.
40
1 0 0 1 0 0
Q - v a lu e
R Q D , %
8 0 1 0
4
6 0
1 .0
4 0
0 .4
2 0
0 0 .1
0 3 0 0 0 3 5 0 0 4 0 0 0 4 5 0 0 5 0 0 0 5 5 0 0
S e is m ic v e lo c ity ( V p ) , m /s
Figure 2.9 Seismic velocity versus RQD value and Q value. After Sjøgren, B.,
Øvsthus, A. And Sandberg, B.: Seismic classification of rock mass quali-
ties, in Geophysical Prospecting 27, and Barton, N. and Grimstad, E.:
The Q-System Following Twenty Years of Application in NMT Support
Selection, in Felsbau vol. 12 no. 6 1994.
2.17 Experience from Back-mapping
Amphibolite
Basalt
Gabbro
Gneiss
Granite
Granitic Gneiss
Limestone
Mica Gneiss
Mica Schist
Phyllite
Quartzite
0 0-I I I-II II II-III

III III-IV
few observations Fissure Class

frequently observed
Figure 2.10 Recorded degree of fracturing for some rock types.
41
Figure 2.10 shows a summary of recorded degree of fracturing based on back-

mapping of bored tunnels. The figure shows that each rock type has a substantial
variation in the Fracture Class, which implies that the degree of fracturing must be
mapped during the site investigations to reduce the uncertainty and the risk of the
project.
2.18 Rock Stress
It has been observed in some cases that the rock stress has a positive influence on the
net penetration rate. The observations have been made in rock stress conditions with
high stress anisotropy and the major principal stress more or less parallel to the tunnel
face. The increased net penetration rate was especially distinctive after a standstill
(e.g. over the weekend), and would last for a few metres.
Figure 2.11 Cross section of a chip from a tunnel exposed to high and anisotropic
rock stress. The chip is from a tunnel bored in similar rock stress condi-
tions as outlined in Figure 2.12.
The influence of a favourable rock stress situation with regard to the net penetration
rate may be explained in two ways:
42
• The high rock stress more or less parallel to the tunnel face helps the propagation
of the chipping cracks. This effect is expected to be more or less instantaneous in
the tunnel section exposed to the high and anisotropic rock stress.
• During a standstill, the high rock stress more or less parallel to the tunnel face will
create fissures by stress relief in front of the tunnel face. This effect is expected to
last for a few metres only.
Figure 2.11 shows an almost rectangular cross section of the chip, compared to the
usually more elliptic cross section of chips, see Figure 1.3 and 1.4 of Project report
1F-98 HARD ROCK TUNNEL BORING The Boring Process. The square cross
section indicates that the high stress anisotropy has influenced the chipping process.
The effect of a favourable rock stress situation on the penetration rate may be simu-
lated through the introduction of a virtual Fracture Class with an orientation of 90° to
the tunnel axis.
ls tre s s
s 1
s 3
Figure 2.12 Possible fracturing pattern ahead of a tunnel exposed to high and aniso-
tropic rock stress. lstress will depend on tunnel diameter, rock stress level,
degree of stress anisotropy and time of TBM standstill.
43
The positive effect of rock stress should not be overestimated. The net penetration
rate may increase, but the machine utilisation may decrease due to the need of extra
rock support work, gripper problems and similar.
The rock stress situation may also be unfavourable. In a more or less lithostatic stress
situation, it is possible that the confining stress will increase the apparent rock
strength and also reduce the influence of the rock mass fracturing due to closing of
fractures.
The field data on the influence of rock stress on the net penetration rate are few. A
possible effect of the rock stress must be carefully evaluated before it is incorporated
in the performance and cost estimations.
44
2. SITE INVESTIGATION 2.2 Rock Drillability
2.2 ROCK DRILLABILITY
2.20 Introduction
The rock drillability affects the net penetration rate and the cutter life. As stated in
Section 2.10, the rock drillability is of less importance for the penetration rate (and
thereby the excavation costs as well) than the rock mass degree of fracturing.
It is recommended to use the DRI, CLI and the rock quartz content as drillability pa-
rameters for tunnel boring. The total mineral composition of the rock is also of im-
portance when evaluating the drillability. The compressive strength, the point load
strength and the Brazilian tensile strength are frequently used to describe the proper-
ties of the intact rock. The relations between the DRI and CLI, and other parameters
are discussed briefly below.
2.21 Rock Sampling
The most important quality of a rock sample is that it is valid for the rock type and
geological zone it is meant to represent. Therefore, an engineering geological model
of the complete tunnel should be established before the rock sample locations are de-
cided.
The sampling programme must acknowledge that even a very large sampling scheme
will only provide an estimate of the rock properties along the given tunnel. The num-
ber of samples for a project should be determined by the variation of the rock proper-
ties and the rock types at the site. Samples should be taken for each rock type or geo-
logical zone, and one sample should not represent more than 1 - 3 km of tunnel. The
samples should be at least 10 kg, and may consist of several blocks or core pieces
with a minimum weight per block of 0.5 kg. Less sample weight may be used, but that
will result in fewer parallel brittleness tests and a larger uncertainty in the indices.
One sample should consist of one rock type only. Mixed face conditions should not be
simulated by composing the sample of two or more rock types. This will only lead to
ambiguous results of the laboratory testing.
45
When it is difficult or very expensive to provide sufficient test material for standard
DRI testing, the Mini-DRI test is an option. In most cases, 500 grams of sample mate-
rial is sufficient for this test.
Weathering will influence the parameters tested in the laboratory by increasing the
DRI value and the CLI value. Of the parameters tested in the laboratory, the SJ value
is most sensitive to weathering. If weathering is observed in the rock sample, Table
2.3 may be used to adjust the tested parameters before DRI and CLI are assessed.
1
Description of Weathering S20 SJ AVS
Weathering is observed 0.95 0.75 1

The sample is clearly weathered 0.9 0.4 1
The sample is weathered
Do not use as sample.
through
Table 2.3 Adjustment factors for weathered rock samples (for rough estimates
only).
1
Depending on the tested rock type. Weathering may cause the strong mineral grains like
quartz to be overexposed in the test and in fact result in a higher AVS value than for non-
weathered samples. Usually weathering will cause the AVS to be less than for a non-
weathered sample. The experience data are not consistent and it is not recommended to
adjust the AVS of weathered samples.
2.22 Data from the Drillability Database
The Figures 2.13 and 2.14 show the variation in the drillability indices for some rock
types. For a more detailed study, se the Project Report 13C-98 DRILLABILITY
Statistics of Drillability Test Results.
The main tendencies shown in the Figures 2.13 and 2.14 are:
• All rock types have a considerable variation in the drillability indices. Hence, the
rock name is not a sufficient description of rock drillability.
• Marble shows good drillability in general, DRI > 60 for all samples.
• Quartzite shows poor drillability in general, DRI < 37 for 75 % of the samples.
46
• Limestone and marble will give high cutter life in general, CLI > 40 for all sam-
ples.
• Quartzite will give very low cutter life in general, CLI < 6 for all samples.
• Granite will give low cutter life in general, CLI < 9 for all samples.
Amphibolite
Basalt
Diorite
Gabbro
Gneiss
Granite
Granitic Gneiss
Greenstone
Limestone
Marble
Mica Gneiss
Mica Schist
Phyllite
Quartzite
Sandstone
Shale
0 10 20 30 40 50 60 70 80 90
Drilling Rate Index, DRI
10% 25% 50% 75% 90% percentiles
Figure 2.13 Recorded Drilling Rate Index for some rock types. Data from Project
Report 13C-98 DRILLABILITY Statistics of Drillability Test Results.
47
Amphibolite
Basalt
Diorite
Gabbro
Gneiss
Granite
Granitic Gneiss
Greenstone
Limestone
Marble
Mica Gneiss
Mica Schist
Phyllite
Quartzite
Sandstone
Shale
0 10 20 30 40 50 60 70 80 90
Cutter Life Index, CLI
10% 25% 50% 75% 90% percentiles
Figure 2.14 Recorded Cutter Life Index for some rock types. Data from Project Re-
port 13C-98 DRILLABILITY Statistics of Drillability Test Results.
2.23 The Compressive Strength
The compressive strength σc is often used as a parameter for rock characterisation. In

Figure 2.15, 80 parallel tests of compressive strength and DRI are compared and
grouped according to rock type. The compressive strength was measured at the De-
partment of Geology and Mineral Resources Engineering at NTNU, following the
ISRM standard. 32 mm diameter cores have been used.
Figure 2.15 shows that the compressive strength may overestimate the drillability of
foliated or schistose rocks such as phyllite, mica gneiss, mica schist, shale and green-
schist. For these rock types the compressive strength is relatively low even for low
DRI values. Furthermore, the variation in the compressive strength is less than the
variation in the DRI.
The risk of underestimation of the variation in the drillability seems to apply to cal-
citic rocks like limestone and marble as well.
48
Sedimentary rocks like siltstone and sandstone, and igneous rocks like fine and coarse
grained granite seem to have a good correlation between DRI and σc.
Amphibolitic gneiss seems to have a larger variation in σc than in DRI.
The compressive strength is quite sensitive to the testing conditions2. Different labo-
ratories measure differing values for the compressive strength for parallel tests of the
same rock samples. Registered deviations have been considerable. The deviations are
caused by differences in the size of samples, water content at testing and applied
loading rate.
C o m p r e s s iv e s tr e n g th , M P a
3 0 0 3 0 0
A m p h ib o litic g n e is s
2 0 0 2 0 0
G re e n s to n e
1 0 0 1 0 0
L im e s to n e M a r b le
G r e e n s c h is t
C a lc a r e o u s s h a le
2 0 3 0 4 0 5 0 6 0 7 0 8 0 9 0 1 0 0 2 0 3 0 4 0 5 0 6 0 7 0 8 0 9 0 1 0 0
D r illin g R a te In d e x , D R I D r illin g R a te In d e x , D R I
Q u a r tz ite
3 0 0 3 0 0
C o a rs e
2 0 0 g r a in e d 2 0 0
g r a n ite
M ic a M e d iu m to fin e
g n e is s g r a in e d g r a n ite
1 0 0 1 0 0 S a n d s to n e
P h y llite M ic a s c h is t
S h a le S ilts to n e
2 0 3 0 4 0 5 0 6 0 7 0 8 0 9 0 1 0 0 2 0 3 0 4 0 5 0 6 0 7 0 8 0 9 0 1 0 0
D r illin g R a te In d e x , D R I D r illin g R a te In d e x , D R I
Figure 2.15 Drilling Rate Index DRI versus compressive strength σc grouped ac-
cording to rock type.
2
Myrvang, A.: Lecture notes in Rock Mechanics, basic course, Trondheim 1996.
49
The compressive or tensile strength of anisotropic rocks depends on the testing direc-
tion versus the planar structure of the rock, while the DRI is not influenced by the
testing direction.
When using σc to estimate the DRI of a rock type, one must ensure that the compres-
sive strength is measured in the laboratory, and not derived from point load strength
measurements. If the latter is done, one must estimate the DRI directly from the point
load index, see Section 2.24.
2.24 The Point Load Strength
E. Broch and J. A. Franklin presented the Point Load Index in 19723, based on the use
of a portable test apparatus. The point load strength test itself has been in use since
before 1960. It is an easy and cheap test to perform. The portable equipment makes it
possible to perform the test in the field. The point load strength and the Brazilian ten-
sile strength are frequently used as parameters for rock characterisation.
P
Is = (MPa) [2.15]
D2
Is = The Point Load Strength Index (MPa)

P = Failure load (N)
D = Distance between the platen points, i.e. core diameter (mm)
The Is for 50 mm diameter cores is selected as reference diameter when comparing

test results from different core diameters. The test results based on other core diame-
ters are transformed according to a correlation chart3.
The point load test may also use test specimens of irregular shape. Hand specimens
should have longest to shortest axis ratio of 1.0 - 1.4 and be tested along the longest
axis. The test results must be corrected for shape and size3.
3
Broch, E., Franklin, J. A.: The Point Load Strength Test in Int. J. Rock Mech. Min. Sci. Vol. 9, 1972.
50
The crushing process in the Brittleness test is believed to mainly consist of tensile
failure. Hence, one may expect some correlation between the DRI value and the point
load strength. Figure 2.16 shows a rough relation based on 27 parallel samples. The
correlation must be used with caution.
D r illin g R a te In d e x , D R I
1 0 0
8 0
6 0
4 0
2 0
5 1 0 1 5 2 0 2 5
P o in t L o a d S tr e n g th , M P a
Figure 2.16 Relation between the point load strength for 25 - 27 mm cores and the
DRI value. After O. T. Blindheim "The Drillability of Rocks", 1979.
51
2.25 CERCHAR Abrasivity Index CAI
The CAI was introduced by the Centre d'Etudes et Recherches des Charbonnages de
France in 1971. The CERCHAR Abrasivity Index may be used for estimation of the
cutter life.
The CERCHAR scratch test uses a steel pin against the rock surface, is easy to per-
form and requires only a small sample specimen. Establishing a correlation between
CAI and CLI is desirable and may increase the use of the prediction models in the
Project Report 1B-98 HARD ROCK TUNNEL BORING Advance Rate and
Cutter Wear.
Figure 2.17 shows a rough correlation of CAI and CLI. The correlation should only be
used for rough estimates.
C A I
6 .0
4 .0
2 .0
1 0 2 0 3 0 4 0 5 0 6 0
C L I
Figure 2.17 Rough correlation of CAI and CLI.
52
2.26 Vickers Hardness Number Rock VHNR
L im e s to n e
7 0 0
V e s ic u la r b a s a lt
6 0 0
(h r) o
5 0 0
c u tte rs , H
P h y llite s
4 0 0
C u tte r r in g life fo r 3 9 4 m m
G r e e n s to n e /g r e e n s c h is t
3 0 0
2 0 0
A r k o s ite
1 0 0
9 0
8 0
M ic a g n e is s
7 0
6 0
G r a n ite /
g n e is s
5 0
M ic a s c h is t/
m ic a g n e is s
4 0
Q u a r tz ite
3 0
M A = % m ic a + % a m p h ib o le
2 0 M A < 1 5 %
1 5 % < M A < 3 5 %
3 5 % < M A < 4 5 %
M A > 4 5 %
1 0
1 0 0 2 0 0 3 0 0 4 0 0 5 0 0 6 0 0 8 0 0 1 0 0 0
T o ta l V ic k e r s H a r d n e s s , V H N R
Figure 2.18 Registered cutter ring life as a function of Vickers hardness number for
rock VHNR. RPM = 38/dtbm and dc = 394 mm.
53
The Vickers hardness number VHNR for a rock type is found by weighting the Vick-
ers hardness of each mineral to a compound Vickers hardness for the rock type (see
the Project Report 13A-98 DRILLABILITY Test Methods).
The VHNR is based on the rock mineral content, which is easy to find by laboratory
testing and requires only a small sample. The rock mineral content may also be de-
rived from general knowledge of the geology without specific laboratory testing.
The Figure 2.18 shows envelope curves for registered and normalised cutter ring life.
The relations are based on relatively few observations, and should be used with cau-
tion.
54
2. SITE INVESTIGATION 2.3 Rock Support
2.3 ROCK SUPPORT
The site investigations should provide a basis to estimate the types and quantities of
rock support. In order to reduce the risk, the TBM and the backup system must be
equipped to handle the expected stability problems and install the necessary rock sup-
port. At the same time the system should not be over-equipped, which would be a
solution that most probably would increase the construction time and costs unneces-
sarily.
In hard rock conditions there are several situations related to the geology that may
cause long delays and extra costs. Field experience indicates that most of those situa-
tions have one common feature: They are singular events that occur with little or no
warning seen from the point of view of the tunnelling crew.
As stated earlier, one must accept a certain risk with regard to the rock conditions
along the tunnel. Normally, the site investigations will reveal rock conditions that
need systematic rock support over longer sections. In addition, the site investigations
must focus on such possible singular events as:
• Major faults or crushed zones.

• Local areas of weak rock.
• Water inflows.
• Locally high rock stress.
A rock support program based on the site investigations must be established and in-
corporated in the design of the machine and the backup system, as well as in the time
and cost estimates.
Time and cost estimates for rock support work are found in the following project re-
ports (at present available in Norwegian editions only):
• 2F-99 TUNNELLING Time Planning

• 10A-91 TUNNEL ROCK SUPPORT Bolts
• 10B-91 TUNNEL ROCK SUPPORT Shotcrete
• 10C-91 TUNNEL ROCK SUPPORT Concrete Lining.
55
2. SITE INVESTIGATION 2.4 Engineering Geological Summary
2.4 ENGINEERING GEOLOGICAL SUMMARY
The results of the site investigations may be presented in an engineering geological

summary. The main purpose of the summary is to help the communication between
all parts involved in the project from planning through bidding to construction.
Figure 2.19 shows an example giving a rough project overview.
56
2. SITE INVESTIGATION 2.4 Engineering Geological Summary
G e o lo g y
2 5 0 0 m 8 0 0 m 3 2 0 0 m
In fo r m a tio n o n r o c k ty p e , s tr e n g th , d r illa b ility in d ic e s , d e g r e e o f fr a c tu r in g , c r u s h e d

z o n e s , w a te r , h is to r y e tc ., g iv e n in m a p s , p r o file s , b o r e h o le lo g g in g a n d r e p o r ts
G r a n itic A m p h ib o litic
R o c k T y p e g n e is s g n e is s G n e is s
R e p re s e n ta tiv e
d r illa b ility in d ic e s D R I= 4 8 D R I= 3 3 D R I= 6 0
( E s tim a te d o n th e C L I= 8 C L I= 1 2 C L I= 2 0
b a s is o f s a m p le s ) Q = 3 0 % Q = 8 % Q = 2 5 %
1 6 0 c m 4 0 c m 2 0 c m
S p a c in g , ty p e n o n - c o n tin u o u s n o n - c o n tin u o u s n o n - c o n tin u o u s
O r ie n ta tio n V a r ie s , 1 0 ° 2 5 ° 4 5 °
F r a c tu r e C la s s o f r o c k S t 0 -I S t I S t II
F o r 4 .5 m d tb m , 4 8 3 m m c u tte rs , 1 1 .1 rp m a n d 1 7 2 0 k W in s ta lle d p o w e r
T h r u s t le v e l 2 7 0 k N /c 2 7 0 k N /c 2 7 0 k N /c
N e t p e n e tr a tio n r a te 2 .0 0 m /h 2 .7 3 m /h 5 .5 3 m /h
C u tte r life 1 .9 9 h /c 1 .4 7 h /c 3 .4 4 h /c
M a c h in e u tilis a tio n 4 7 .0 % 4 0 .6 % 4 0 .5 %
W e e k ly a d v a n c e fo r 9 5 m 1 1 2 m 2 2 6 m
1 0 1 h /w e e k
S m a lle r le a k s a r e e x p e c te d fr o m s in g le jo in ts in g r a n itic g n e is s ,
W a te r
i.e . tu n n e l s h o u ld b e e x c a v a te d o n in c lin a tio n
R o c k s u p p o rt
N o r o c k s u p p o r t o f s ig n ific ia n c e e x p e c te d , o c c a s io n a l s c a lin g a n d
M a jo r s e c tio n s
fe w b o lts
S h o r t s e c tio n s 2 5 0 m
E s tim a te d le n g th R o c k fa lls fr o m tu n n e l r o o f in
o f r o c k s p a llin g c r u s h e d z o n e s , p lu s g r ip p e r
p r o b le m s
T y p e S y s te m a tic b o ltin g B o ltin g a n d s tr a p s o r

in r o o f, p o s s ib ly fib r e - r e in fo r c e d s h o tc r e te
c o m b in e d w ith s tr a p s
A d d itio n a l tim e
2 w e e k s 5 d a y s 3 d a y s 2 d a y s
R o c k s u p p o rt
O c c a s io n a l s c a lin g a n d b o ltin g 6 h /k m
U n fo re s e e n ro c k
c o n d itio n s a n d m a jo r 2 w e e k s 1 w e e k 3 w e e k s
m a c h in e s to p s
Figure 2.19 Summary of engineering geological information.

57
2. SITE INVESTIGATION 2.5 Variability of the Rock Conditions
2.5 VARIABILITY OF THE ROCK CONDITIONS
The site investigations are not able to completely uncover the rock conditions along
the tunnel and hard rock tunnel boring is very sensitive to changes in the rock condi-
tions. As a supplement to the time and cost estimates for the expected geological con-
ditions, one should also carry out an assessment of the influence on the time and cost
estimates of changed rock conditions. This may be done for one positive and one
negative variant with regard to the rock mass boreability. The positive or negative
variant must not be interpreted as the best or worst case, but rather as rock mass con-
ditions that describe a "natural" or commonly accepted variation in the rock mass
properties.
Such an evaluation is of course highly subjective, but some help may be found in the
Project Report 13C-98 DRILLABILITY Statistics of Drillability Test Results.
The report contains cumulative distributions of drillability indices for a number of
rock types, see example in Figure 2.20. The figure also shows examples of index val-
ues describing the variation in the rock mass properties.
When the positive and negative variants of the rock conditions are modelled, there are
some important points to be considered.
• If the index value related to the expected rock conditions is in the middle of the
cumulative distribution, one may use an equal offset in cumulative percentage up
and down for the positive and negative variants, e.g. 15 %.
• If the index value related to the expected rock conditions is in the upper or lower
part of the cumulative distribution, one should use different offsets up and down
for the positive and negative variants. One may also use less offsets than for val-
ues in the middle of the distribution. E.g. if the expected value corresponds to a
cumulative percentage of 25, one may use offsets of 10 % upwards and 7.5 %
downwards.
Variation in the rock mass degree of fracturing may be roughly estimated from Figure
2.10 in Section 2.17. It is recommended that the variation in the degree of fracturing
is selected as approximately one half Fracture Class up and down from the expected
conditions.
58
B O R B A R d a ta b a s e
N o r m a l d is tr ib u tio n
8 0
C u m u la tiv e p e r c e n ta g e
6 5 %
6 0
5 0 %
4 0
3 5 %
2 5 %
2 0 1 7 .5 %
3 6 4 0 4 6 5 1
3 2
2 0 4 0 6 0 8 0
D R I
Figure 2.20 Cumulative distribution of DRI for sandstone.
Rock Conditions Negative Expected Positive

Degree of fracturing St I St I-II St II
Angle α 15° 20° 20°
DRI 41 46 52
CLI 6.1 7.7 10.6
Quartz content 30 % 25 % 25 %
Table 2.4 Variants of an expected data set for sandstone.
There are several tools available to help estimate the uncertainty and risk of construc-
tion projects. General tools developed at the Department of Building and Construction
Engineering are well suited to use in tunnel projects, and include:
59
• TIDUS - a PC program for time scheduling under uncertainty, based on the suc-
cessive principle
• TRIKALK - a PC program for cost estimation under uncertainty, based on the
successive principle
• RISIKINI - a PC programme for risk analysis based on binary event trees.
60
APPENDIX A. Previous Editions
A. PREVIOUS EDITIONS
Previous editions of the Hard Rock Tunnel Boring Report including project group
members:
1-76 Norwegian edition

Bengt Drageset
Roy-Egil Hovde
Erik Dahl Johansen
Roar Sandnes
O. Torgeir Blindheim
Odd Johannessen
1-79 Norwegian edition

Knut Gakkestad
Jan Helgebostad
Svein Paulsen
Oddbjørn Aasen
Erik Dahl Johansen
O. Torgeir Blindheim
Odd Johannessen
1-83 Norwegian and English edition

Arne Lislerud
Steinar Johannessen
Amund Bruland
Tore Movinkel
Odd Johannessen

Arne Lislerud
Amund Bruland
Bjørn-Erik Johannessen
Tore Movinkel
Karsten Myrvold
Odd Johannessen

Bård Sandberg
Amund Bruland
Jan Lima
Odd Johannessen
61
APPENDIX B. Research Partners
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
62
APPENDIX C. List of Parameters
C. List of Parameters
The parameters used in the report are listed in the following. The list is according to
when the parameter first is explained or treated.
Parameter Description Unit Page
af Average spacing between planes of weakness cm 14

AV Abrasion Value (tungsten carbide) mg/5 min 9
(see also PR 13A-98)
AVS Abrasion Value Steel (cutter ring steel) mg/1 min 9
CAI CERCHAR Abrasivity Index 52
CLI Cutter Life Index 9
dc Cutter diameter mm 53
(see also PR 1B-98)
dtbm TBM or cutterhead diameter m 53
(see also PR 1B-98)
D Distance between the platen points in
the point load test mm 50
DRI Drilling Rate Index 7
ESP Type of fracturing - See MSJ 14
Is Point load strength index MPa 50
Ja Parameter related to the joint wall alteration
- used in the Q-System 35
Jn Parameter related to the number of joint sets
Jr Parameter related to the roughness of joints
Jw Parameter related to the water characteristics of
the rock mass - used in the Q-System 35
kQ Correction factor for rock quartz content
- cutter ring life 10
ksi Fracturing factor for set no. i (see also PR 1B-98) 16
ks-tot Total fracturing factor (see also PR 1B-98) 16
lc length of core section - used in core logging m 33
63
lstress Tunnel length influenced by stress cracks m 43

MSJ Type of fracturing - Marked Single Joint 14
nf Number of fractures of a set in a core section
- used in core logging 33
P Failure load in the point load test N 50
Q Rock mass Q-value according to the Q-System 35
Q Rock quartz content % 10
RMR Rock Mass Rating or the Geomechanics
Classification of a rock mass 37
RMRadj Adjusted RMR value - adjusted for orientation
of discontinuities 37
RMRbasic Basic RMR value - not adjusted for orientation
of discontinuities 37
RMRi Rating of the individual parameters
in the RMR System 38
RPM Cutterhead revolutions per minute rev/min 53
(see also PR 1B-98)
RQD Rock Quality Designation % 33
S20 Brittleness Value after 20 impacts % 7
SJ Sievers' J-value (miniature drill test) mm/10 8
Sp Type of fracturing - joints 13
SRF The stress condition of the rock mass
St Type of fracturing - fissures 13
Vp Longitudinal seismic velocity in rock m/s 41
VHNR Vickers Hardness Number Rock 53
α Angle between the tunnel axis and the planes of
weakness (see also PR 1B-98) ° 59
αc Angle of intersection between a fracture set and
g
the core axis - used in core logging or ° 33
αd Dip direction ° or g 29
64
αdn Dip direction of the normal to a discontinuity plane ° or g 30

αf Dip angle ° or g 29
αfn Dip angle of the normal to a discontinuity plane ° or g 30
αs Strike angle ° or g 29
αsn Strike direction of the normal to a discontinuity plane ° or g 30
αt Direction of the tunnel axis ° or g 32
σc Unconfined or uniaxial compressive strength MPa 48
σ1 Maximum principal rock stress MPa 43
σ3 Minimum principal rock stress MPa 43
65
APPENDIX D. The Q-System
D. THE Q-SYSTEM
N. Barton, R. Lien and J. Lunde at the Norwegian Geotechnical Institute presented

the Q-system in 1974. The method has been revised in 1993. The revised Q-system is
described in: Barton, Nick and Grimstad, Eystein 1994: The Q-System following
Twenty Years of Application in NMT Support Selection. Felsbau 12, No. 6, pp 428 -
436.
The Q value of a rock mass is found by combining six parameters.
RQD J r J w
Q= ⋅ ⋅ [D.1]
J n J a SRF
RQD = Rock Quality Designation or degree of jointing

Jn = relates to the number of joint sets
Jr = relates to the roughness of the most important joints
Ja = relates to the joint wall rock alteration and/or the joint filling material
Jw = relates to the joint water leakage or pressure
SRF = relates to the stress condition in the rock mass.
RQD
is a measure of block size
Jn
Jr
is a measure of interblock friction angle
Ja
Jw
is a measure of the active stress
SRF
1. Rock Quality Designation RQD

A Very poor 0 - 25
B Poor 25 - 50
C Fair 50 - 75
D Good 75 - 90
E Excellent 90 - 100
Note: i) Where RQD is reported or measured as ≤ 10 (including 0), a nominal value of 10 is
used to evaluate Q.
ii) RQD intervals of 5, i.e. 100, 95, 90, etc., are sufficiently accurate.
66
2. Joint Set Number Jn

A Massive, no or few joints 0.5 - 1.0
B One joint set 2
C One joint set plus random joints 3
D Two joint sets 4
E Two joint sets plus random joints 6
F Three joint sets 9
G Three joint sets plus random joints 12
Four or more joint sets, random, heavily jointed,
H 15
"sugar cube", etc.
J Crushed rock, earthlike 20
Note: i) For intersections, use 3.0 ⋅ Jn
ii) For portals, use 2.0 ⋅ Jn
3. Joint Roughness Number Jr

a) Rock - wall contact, and b) rock - wall contact before 10 cm shear
A Discontinuous joints 4
B Rough or irregular, undulating 3
C Smooth, undulating 2
D Slickensided, undulating 1.5
E Rough or irregular, planar 1.5
F Smooth, planar 1.0
G Slickensided, planar 0.5
Note: i) Descriptions refer to small scale features and intermediate scale features,
in that order
c) No rock - wall contact when sheared

Zone containing clay minerals thick enough to prevent
H 1.0
rock - wall contact
Sandy, gravelly or crushed zone thick enough to prevent
J 1.0
rock - wall contact
Note: i) Add 1.0 if the mean spacing of the relevant joint set is greater than 3 m.
ii) Jr = 0.5 can be used for planar slickensided joints having lineations, provided the
lineations are oriented for minimum strength.
67
φr
4. Joint Alteration Number Ja
approx.
a) Rock - wall contact (no mineral fillings, only coatings)
Tightly healed, hard, non-softening,
A - 0.75
impermeable filling, i.e. quartz or epidote
B Unaltered joint walls, surface staining only 25° - 35° 1.0
Slightly altered joint walls. Non-softening
C mineral coatings, sandy particles, clay-free 25° - 30° 2.0
disintegrated rock, etc.
Silty or sandy clay coatings, small clay
D 20° - 25° 3.0
fraction (non-softening)
Softening or low friction clay mineral
coatings, i.e. kaolinite or mica. Also
E 8° - 16° 4.0
chlorite, talk, gypsum, graphite, etc., and
small quantities of swelling clays.
b) Rock - wall contact before 10 cm shear (thin mineral fillings)
Sandy particles, clay-free disintegrated
F 25° - 30° 4.0
rock, etc.
Strongly over-consolidated non-softening
G clay mineral fillings (continuous, 16° - 24° 6.0
but < 5 mm thickness)
Medium or low over-consolidation,
H softening, clay mineral fillings (continuous, 12° - 16° 8.0
but < 5 mm thickness)
Swelling clay fillings, i.e. montmorillonite
(continuous, but < 5 mm thickness). Value
J 6° - 12° 8 - 12
of Ja depends on percent of swelling clay-
size particles, and access to water, etc.
c) No rock - wall contact when sheared (thick mineral fillings)
Zones or bands of disintegrated or crushed
KL 6, 8 or
rock and clay (see G, H, J for description of 6° - 24°
M 8 - 12
clay condition)
Zones or bands of silty or sandy clay, small
N - 5.0
clay fraction (non-softening)
Thick, continuous zones or bands of clay
OP 10, 13
(see G, H, J for description of clay 6° - 24°
R or 13 - 20
condition)
68
Approx.
5. Joint Water Reduction Factor water pres. Jw
(kg/cm2)
Dry excavations or minor inflow,
A <1 1.0
i.e. < 5 l/min locally
Medium inflow or pressure, occasional
B 1 - 2.5 0.66
outwash of joint fillings
Large inflow or high pressure in competent
C 2.5 - 10 0.5
rock with unfilled joints
Large inflow or high pressure, considerable
D 2.5 - 10 0.33
outwash of joint fillings
Exceptionally high inflow or water pressure
E > 10 0.2 - 0.1
at blasting, decaying with time
Exceptionally high inflow or water pressure
F > 10 0.1 - 0.05
continuing without noticeable decay
Note: i) Factors C to F are crude estimates. Increase Jw if drainage measures are installed.
ii) Special problems caused by ice formation are not considered.
6. Stress Reduction Factor SRF

a) Weakness zones intersecting excavation, which may cause loosening
of rock mass when tunnel is excavated
Multiple occurrences of weakness zones containing clay or
A chemically disintegrated rock, very loose surrounding rock 10
(any depth)
Single weakness zones containing clay or chemically
B 5
disintegrated rock (depth of excavation ≤ 50 m)
Single weakness zones containing clay or chemically
C 2.5
disintegrated rock (depth of excavation > 50 m)
Multiple shear zones in competent rock (clay-free), loose
D 7.5
surrounding rock (any depth)
Single shear zones in competent rock (clay-free) (depth of
E 5
excavation ≤ 50 m)
Single shear zones in competent rock (clay-free) (depth of
F 2.5
excavation > 50 m)
Loose, open joints, heavily jointed or "sugar cube", etc.
G 5
(any depth)
Note: i) Reduce these values of SRF by 25 - 50 % if the relevant shear zones only influence
but do not intersect the excavation.
69
6. Stress Reduction Factor σc /σ1 σθ /σc SRF
b) Competent rock, rock stress problems

H Low stress, near surface, open joints > 200 < 0.01 2.5
J Medium stress, favourable stress condition 200 - 10 0.01 - 0.3 1
High stress, very tight structure. Usually
K favourable to stability, may be unfavourable 10 - 5 0.3 - 0.4 0.5 - 2
for wall stability
Moderate slabbing after > 1 hour in
L 5-3 0.5 - 0.65 5 - 50
massive rock
Slabbing and rock burst after a few minutes
M 3-2 0.65 - 1 50 - 200
in massive rock
Heavy rock burst (strain-burst) and
N immediate dynamic deformation in massive <2 >1 200 - 400
rock
Note: ii) For strongly anisotropic virgin stress field (if measured): when 5 ≤ σ1/σ3 ≤ 10, reduce σc
0.75 ⋅ σc. When σ1/σ3 > 10, reduce σc to 0.5 ⋅ σc , where σc = unconfined compression
strength, σ1 and σ3 are the major and minor principal stresses, and σθ = maximum
tangential stress (estimated from elastic theory).
iii) Few case records available where depth of crown below surface is less than span
width. Suggest SRF increase from 2.5 to 5 for such cases (see H).
6. Stress Reduction Factor σθ /σc SRF
c) Squeezing rock: plastic flow of incompetent rock under the influence of high
rock pressure
O Mild squeezing rock pressure 1-5 5 - 10
P Heavy squeezing rock pressure >5 10 - 20
1/3
Note: iv) Cases of squeezing rock may occur for depth H > 350 ⋅ Q (Singh et al. 1992). Rock
1/3
mass compression strength can be estimated from q = 7 ⋅ γ ⋅ Q (MPa) where γ = rock
3
density in g/cm (Sing 1993).
d) Swelling rock: chemical swelling activity depending on presence of water

R Mild swelling rock pressure 5 - 10
S Heavy swelling rock pressure 10 - 15
Note: Jr and Ja classification is applied to the joint set or discontinuity that is least
favourable for stability both from the point of view of orientation and shear
resistance (where τ ≈ σn ⋅ tan-1(Jr /Ja)).
70
ROCK MASS CLASSIFICATION

Q Rock Classes
0.001 - 0.01 G Exceptionally poor
0.01 - 0.1 F Extremely poor
0.1 - 1 E Very poor
1-4 D Poor
4 - 10 C Fair
10 - 40 B Good
40 - 100 Very good
100 - 400 A Extremely good
400 - 1000 Exceptionally good
71
APPENDIX E. The RMR System
E. THE RMR SYSTEM1
Z. T. Bieniawski presented the Rock Mass Rating System or the Geomechanics Clas-
sification in 1973. The system was modified over the years until 19792.
According to Bieniawski2, the first step is to divide the rock mass "into structural re-
gions such that certain features are more or less uniform within each region".
The basic RMR value varies from 0 to 100 and is found by summing the ratings of
five rock mass parameters.
RMRbasic = RMR1 + RMR2 + RMR3 + RMR4 + RMR5 [E.1]
RMR1 = The strength of the intact rock material.

RMR2 = RQD Rock Quality Designation.
RMR3 = The average spacing of discontinuities.
RMR4 = The condition of discontinuities.
RMR5 = The groundwater conditions.
In addition, the RMR value may be adjusted for the type of excavation and the orien-
tation of discontinuities relative to the excavation (RMR6).
RMRadj = RMRbasic + RMR6 [E.2]
The 19792 and 19891 descriptions of the Geomechanics Classification are slightly
different in the description of how to evaluate the rating of each parameter. In 1979 it
is said that average conditions are evaluated for the basic RMR value, and that the
ratings given for discontinuity spacings apply to rock masses having three sets of dis-
continuities.
The 1989 edition indicates that the RMR value of the rock mass may be estimated
through two approaches:
1
Z. T. Bieniawski 1989: Engineering Rock Mass Classifications, John Wiley & Sons.
2
Z. T. Bieniawski 1979: The Geomechanics Classification in Rock Engineering Applications. Proceedings
of the 4th International Congress on Rock Mechanics, ISRM, Montreux, Vol. 2, pp 41 - 48
72
• If the type and orientation of the construction or excavation is known, the ratings
of the discontinuity set that will control the stability of the rock mass are summed
and will constitute the overall RMR.
• Where no set of discontinuities is dominant or of critical importance, or when es-
timating the general rock mass strength and deformability, the ratings of each dis-
continuity set are averaged for each individual classification parameter and then
summed to constitute the RMR.
For more details on the use and extensions of the RMR System, we refer to various
publications concerning the Geomechanics Classification, of which Z. T. Bienieawski
has published several.
Strength of Intact Rock Material

Point Load Strength Uniaxial Compressive
OR RMR1
Index (MPa) Strength (MPa)
> 10 > 250 15
4 - 10 100 - 250 12
2-4 50 - 100 7
1-2 25 - 50 4
Do not use, UCS is preferred 5 - 25 2
Do not use, UCS is preferred 1-5 1
Do not use, UCS is preferred <1 0
Drill Core Quality RQD (%) RMR2

90 - 100 20
75 - 90 17
50 - 75 13
25 - 50 8
< 25 3
73
Spacing of Discontinuities (m) RMR3

> 2.0 20
0.6 - 2.0 15
0.2 - 0.6 10
0.06 - 0.2 8
< 0.06 5
Condition of Discontinuities RMR4

Very rough surfaces. Not continuous. No separation.
30
Unweathered wall rock.
Slightly rough surfaces. Separation < 1 mm.
25
Slightly weathered walls.
Slightly rough surfaces: Separation < 1 mm.
20
Highly weathered wall.
Slickensided surfaces OR Gauge < 5 mm thick
OR Separation 1 - 5 mm. 10
Continuous.
Soft gouge > 5 mm thick
OR Separation > 5 mm. 0
Continuous.
Ground Water
Inflow per 10 m Joint water pressure (MPa)
General
Tunnel Length OR Major principal stress (MPa) OR RMR5
Conditions
(l/min)
Completely
None 0 15
dry
< 10 0.0 - 0.1 Damp 10
10 - 25 0.1 - 0.2 Wet 7
25 - 125 0.2 - 0.5 Dripping 4
> 125 > 0.5 Flowing 0
74
Rating Adjustment for Discontinuity Orientations
Strike and Dip Orienta- RMR6

tions of Discontinui-
ties Tunnels Foundations Slopes Excavations1
Very favourable 0 0 0 -12

Favourable -2 -2 -5 -10
Fair -5 -7 -25 -5
Unfavourable -10 -15 -50 -2
Very unfavourable -12 -25 -60 0
1
The ratings for "Excavations" are according to: Fowell, R. J., Johnson, S. T.: Cuttability assessment ap-
plied to drag tool tunnelling machines. Proceedings of the ISRM Congress on Rock Mechanics in Aachen.
Balkema, Rotterdam 1991.
Effect of Discontinuity Strike and Dip Orientations in Tunnelling

Strike Orientation Dip Description
45° - 90° Very favourable

Drive with dip
Perpendicular to 20° - 45° Favourable
tunnel axis 45° - 90° Fair
Drive against dip
20° - 45° Unfavourable
45° - 90° Very unfavourable
Parallel to tunnel axis
20° - 45° Fair
Irrespective of strike 0° - 20° Unfavourable
Rock Mass Classes Determined from Total Ratings

Rating Class No. Description
100 - 81 I Very good rock
80 - 61 II Good rock
60 - 41 III Fair rock
40 - 21 IV Poor rock
≤ 20 V Very poor rock
75
Meaning of Rock Mass Classes

Average Stand-up Cohesion of the Friction Angle of
Class No.
Time Rock Mass (kPa) the Rock Mass (°°)
I 20 years for 15 m span > 400 > 45
II 1 year for 10 m span 300 - 400 35 - 45
III 1 week for 5 m span 200 - 300 25 - 35
IV 10 h for 2-5 m span 100 - 200 15 - 25
V 30 min for 1 m span < 100 < 15
76
APPENDIX F. Mean Orientation of Discontinuities
F. MEAN ORIENTATION OF DISCONTINUITIES
The following describes how to find the mean or preferred orientation of one set of
discontinuities by using the vectorial sum of the normals to the discontinuity planes.
The method is described by R. E. Goodman in Introduction to Rock Mechanics, 2nd
Edition, John Wiley & Sons, 1989.
Each normal is considered as a unit vector and the orientation of the resultant vector
of all individual vectors in a set represents the mean or preferred orientation of the
set. The orientation of the resultant is found by accumulating the direction cosines of
the normals.
a l
N o rm
n
y N o rth
d
l
b P r o je c tio n o f
n o rm a l
m
x E a s t
Figure F.1 Co-ordinate system and projection angles.
77
The measured strike and dip are transformed to the horisontal and vertical projection
angles of the normal according to Figure F.1. β is positive counterclockwise from the
x-axis. For the upper hemisphere δ is positive when the normal rises above the
horisontal xy-plane. For the lower hemisphere δ is positive when the normal points
below the horisontal xy-plane.
The projection angles will depend on which hemisphere one uses for the normals. For
normals in the upper hemisphere, the transformations of angles are:
β = 100 g − (α s + 100 g )= −α s (g ) [F.1]
δ = 100 g − α f (g ) [F.2]
αs = strike angle (0g - 400g)

αf = dip angle (0g - 100g)
For normals in the lower hemisphere, the transformations of angles are:
β = 100 g − (α s + 300 g )= −α s − 200 g (g ) [F.3]
δ = 100 g − α f (g ) [F.4]
The direction cosines are calculated from β and δ according to Figure F.1.
l = cos δ ⋅ cos β [F.5]
m = cos δ ⋅ sin β [F.6]
78
n = sin δ [F.7]
The direction cosines of the resultant vector of the sum of the normals in a set is
found by the sum of all individual direction cosines divided by the magnitude of the
resultant vector as in [F.9] - [F.11]. The magnitude of the resultant vector is found by
[F.8].
R = (∑ li )2 + (∑ mi )2 + (∑ ni )2 [F.8]
lR =
∑ li [F.9]
R
mR =
∑ mi [F.10]
R
nR =
∑ ni [F.11]
R
The projection angles of the resultant vector or mean normal are found from [F.5] -
[F.7] as shown in [F.12] and [F.13].
δ R = arcsin (n R ) 0 ≤ δ R ≤ 100 g (g ) [F.12]
79
 l 
β R = arccos R  mR ≥ 0 (g )
 cos δ R 
 l 
β R = − arccos R  mR < 0 (g ) [F.13]
 cos δ R 
β R = 0g δ R = 100 g
βR and δR are in the range -200g - 200g. When solving [F.1] - [F.4] to find the
resulting strike and dip, one must relate this to the 0g - 400g system in which the strike
and dip measurements have been made. For the upper hemisphere, the mean or
preferred strike and dip are found by [F.14] and [F.15].
α sR = − β R (g ) β R ≤ 0g
[F.14]
α sR = 400 g − β (g ) β R > 0g
α fR = 100 g − δ R (g ) [F.15]
For normals plotted in the lower hemisphere, the mean or preferred strike and dip are
found by [F.16] and [F.17].
α sR = − β R + 200 g (g ) β R ≤ 0g
[F.16]
α sR = 200 g − β R (g ) β R > 0g
α fR = 100 g − δ R (g ) [F.17]
80
ISBN 82-471-0281-1
ISSN 0802-3271

NTNU Hard Rock Tunnel Boring Thesis

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Geology and Site Investigations

Hard Rock Tunnel Boring

0.1 Project Reports about Hard Rock Tunnel Boring 3

1 ROCK MASS BOREABILITY 6

1.1 Rock Drillability 7

1.2 Rock Mass Fractures 12

2.1 Degree of Fracturing 27

2.2 Rock Drillability 45

2.3 Rock Support 55

2.4 Engineering Geological Summary 56

2.5 Variability of the Rock Conditions 58

E. The RMR System 72

F. Mean Orientation of Discontinuities 77

HARD ROCK TUNNEL BORING Geology and Site Investigations

• 1A-98 HARD ROCK TUNNEL BORING Design and Construction

In addition, HARD ROCK TUNNEL BORING Background and Discussion gives

For reference, registration and similar, we ask for the following:

NTNU-Anleggsdrift (1998): Project Report 1D-98 HARD ROCK TUNNEL

When copying from the report, the source should be stated.

Trondheim, April 2000

Contact address: Amund Bruland

Telephone +47 73 59 47 37 Fax +47 73 59 70 21

0.1 PROJECT REPORTS ABOUT HARD ROCK TUNNEL BORING

The report treats the following items:

Appendix A shows a list of previous editions of the HARD ROCK TUNNEL

Use of the Estimation Models

• Preliminary and feasibility studies

• Estimating net penetration rate and cutter wear

• The rock mass type of fracturing

1.1 ROCK DRILLABILITY

1.12 The Drilling Rate Index DRI

1.13 Rock Abrasiveness

• The Abrasion Value AV using tungsten carbide as test piece.

1.14 The Cutter Life Index CLI

1.15 Rock Mineral Content

1.2 ROCK MASS FRACTURES

Planes of weakness or discontinuities in the rock mass contribute considerably to the

Rock Mass Properties Machine Factors

• Type, frequency and orientation of the • Thrust level

1.21 Classification of Fractures

• Systematically fractured rock mass.

The classification of weakness planes includes:

• Type of planes with regard to persistence, openness or aperture, clay or mineral

• Fissures (denoted St): Includes non-continuous joints or fissures (can only be

• Marked Single Joints (denoted ESP or MSJ): Includes marked discontinuities

1.22 Degree of Fracturing

Fracture Class Spacing between the Grouping of Classes

Average spacing between the Fissures

ks-tot = total fracturing factor

Figure 1.4 Rock mass with two sets of systematic fractures.

1.23 Examples of Classification

Figure 1.5 Mica gneiss, Fissure Class I-.

Figure 1.7 Granitic gneiss, Joint Class 0-I.

Figure 1.10 Mica schist with Fissure Class IV.

Figure 1.11 Mica gneiss with Fissure Class III+.

Figure 1.12 Mica schist with Fissure Class II.

Figure 1.13 Gneiss with Fissure Class II-.

1.24 Marked Single Joints