Rock Mass Classification and also a brief description of Rock Mass Rating (RMR), Rock Structure Rating (RSR), Q valves and New Austrian Tunneling method (NATM)
Rock Mass Classification and also a brief description of Rock Mass Rating (RMR), Rock Structure Rating (RSR), Q valves and New Austrian Tunneling method (NATM)
The objective of this project is to calculate the factor of safety of a complex slope situation. The stress distribution zones are also shown in the project. The probability of slope failure can be shown using FLAC3D software.
Rocks mechanics and its application in mining geology.
It aims at enhancing the mining process and higher yielding by reducing the chance of failures by providing information about the rocks of the mining area.
Longwall; Longwall in coal; Longwall in Hard Rock; Sublevel Caving; Characteristics of the ore body and mining method; Development; Production; Equipments Used; Block Caving, Introduction, Historical evolution of the method, Condition deposit; Principles of the method; Methodology of block caving; Basic issues of geomechanical to the black caving method; Caveability;Mine design Block caving; Fragmentation and extraction control; Subsidence associated; Advantages and Disadvantages of Block Caving
The objective of this project is to calculate the factor of safety of a complex slope situation. The stress distribution zones are also shown in the project. The probability of slope failure can be shown using FLAC3D software.
Rocks mechanics and its application in mining geology.
It aims at enhancing the mining process and higher yielding by reducing the chance of failures by providing information about the rocks of the mining area.
Longwall; Longwall in coal; Longwall in Hard Rock; Sublevel Caving; Characteristics of the ore body and mining method; Development; Production; Equipments Used; Block Caving, Introduction, Historical evolution of the method, Condition deposit; Principles of the method; Methodology of block caving; Basic issues of geomechanical to the black caving method; Caveability;Mine design Block caving; Fragmentation and extraction control; Subsidence associated; Advantages and Disadvantages of Block Caving
Techniques for measuring insitu stressesZeeshan Afzal
There are some methods that tells about insitu stresses and these are very important methods in Geology as well as well coring and also digging of well as well as in mining these methods are very helpful. So, main idea about is to information about these methods.
insitu stress field in earth crust, stress environment in mines, effects of horizontal stress, control measures of horizontal stress, stress mapping, measurement of insitu stress field
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Back analysis of high tunnel convergences in clayey marlsSYSTRA
Ganntas Tunnel is part of the modernization project of the
railway between Alger and Oran, in Algeria. In order to double
and rectify the existing line between El Affroun and Khémis
Miliana, the alignment foresees the excavation of a 7km-long
twin tunnel. The excavation works started in June 2011 with
the contractor CCECC, under the supervision of SYSTRA.
Excavation is driven in conventional method by hydraulic
hammer simultaneously on 8 different faces since excavation
was started also from a junction window towards the middle
of the tunnel. The minimum longitudinal distance to be
respected between two contiguous tunnel faces has been set
to 30m. The tunnel cross-section is a 70m² oval shaped profile,
temporary support consists of shotcrete, bolts and steel ribs.
A 30 to 50cm thick cast in situ concrete final lining is provided
as well.
When the tunnel reached a fault zone in soft clayey marls,
extreme squeezing occurred, works were stopped, and reprofiling
operations were carried out along more than 100m
tunnel length. To date, works proceed at slow rate since high
convergences are still monitored and completion of works is
not expected before December 2016.
Written by SYSTRA tunnel experts : MISANO Edoardo, COUBRAY Jean-Louis,
ESPINOZA CARMONA Fabiola
A New geotechnical method for natural slope exploration and analysisRasika Athapaththu
Geotechnical investigation of natural slopes is challengeable especially when
natural slopes having higher gradients and access is difficult. Also, it is even more prob-
lematic to find the shear strength parameters spatially to evaluate the stability of slopes as
most of the methods available to find the shear strength parameters in the literature are
uneconomical or such methods cannot be applied in vegetated slopes. Recently, authors
have conducted a series of in situ investigations based on the newly developed lightweight
dynamic cone penetrometer to examine its applicability in analyzing the slopes covered
with weathering remnants of decomposed granite. Six patterns were identified based on the
penetration resistance varies with the depth. Spatial variability analysis conducted on
different grid spaces showed that the coefficient of variation of cone resistance varies from
0 to 35 %. Semi-variogram analysis showed that the Spherical Models can be used to
evaluate the spatial variability of weathering remnants of decomposed granite. A series of
laboratory calibration tests based on the lightweight dynamic cone penetration tests and
direct shear tests with pore pressure measurements were conducted at different void ratios
and degrees of saturation. Based on the laboratory calibration test results, a method to
determine the void ratio, e, from the data of qd was presented. Based on this, two formulas
to evaluate the shear strength parameters, apparent cohesion and friction angle, were
established with the cone resistance and degree of saturation. Slight modification was
proposed in evaluating the apparent cohesion with respect to the different fine content in
the soils. As a whole, the proposed method can be successfully applied to individual slopes
to determine the profile thickness and to evaluate the shear strength parameters spatially.
Based on this, hazard assessment of individual slopes can be made.
1. Keith Kong FICE FIMMM MHKIE CEng RPE(G)
1
Rock Mechanics and Rock Cavern Design
29 November 2016
2. 2
Black & Veatch (HK & SG) involved in
Underground Space Development
3. List of Past 20 years Underground Space Projects
Kau Shat Wan Underground Magazine, (1997)
1.42 km long tunnel (including adits to caverns), Caverns size: 6.5 m span,
5.5 m high, 20m long and rectangular chamber of 13 m wide x 6.8 m high
Tai Po Treatment Works Raw and Treated Water
Aqueducts, (2001)
12 km, 3.8 m dia. of TBM tunnel and 2.5 km drill & blast tunnels with span
3.8 m to 14 m span, 6m height
West Kowloon Drainage Improvement – Tai Hang
Tung Storage Scheme, (2004)
136 m x 130 m and 9.5 m deep storage tank under the existing rugby pitch
and football pitch
Tsim Sha Tsui East Station - Signal Hill Tunnel
(pedestrian subway), (2005)
120 m long, 12 m wide x 9.5 m high horse shoe shaped
Hong Kong West Drainage Tunnel (FS), (2005)
10.5 km with tunnel size 6m and 8m, plus 7.9 km of adits with dia. 2.5 m
& 3.5m dia.
HEC Bowen Road to Kennedy Road Cable Tunnel,
(2008)
0.23 km tunnel, 2.5 m wide x 2.8 m high horseshoe shaped tunnel and
two joint bay caverns 3.3 m wide 4.8m high
Underground Service Reservoir behind The
University of Hong Kong Centennial Campus,
(2009)
Caverns size 15m span & 15m high; tunnel span 8m
Happy Valley Underground Stormwater Storage
Scheme (2015)
The underground storage tank with capacity of 60,000 m³ under the
existing rugby pitch, football pitch and race course.
Sai Kung Sewage Treatment Works to Cavern (FS),
(current)
Process caverns 20m span, 13 – 15m high.
Diamond Hill Service Reservoirs to Cavern (FS),
(current)
Proposed caverns size 18m x 15m
Underground Drainage and Reservoir System,
Singapore (current)
The storage volume for the UDRS is expected to be 100 Mm³
4. 4
Kau Shat Wan Underground Explosives Magazine, 1997
Portals
5. 5
Existing Western Fresh
Water and Salt Water
Pumping Station
New Salt Water
Service Reservoirs
in Caverns
New Pipe Gallery
New Fresh
Water Service
Reservoirs
Historic
Building
Underground Service Reservoir behind HKU
6. 6
17.6 m Span Excavation
(Header)
7 m
(approx.)
Underground Service Reservoir behind HKU
7. 7
Agenda
Ground Investigation and Rock Parameters
In-situ Stress Considerations
Joint Orientations and Effects
Intact Rock and Rock Mass
Rockmass Classifications
Rock Support / Rock Reinforcement Design
Pillar Stability Analysis
8. 8
Rock Mechanics is the subject concerned with the response of
rock to an applied disturbance, which is considered here as an
engineering, i.e. a man-induced disturbance. For a natural
disturbance, rock mechanics would apply to the deformation of
rocks in a structural geology context, i.e. how the folds, faults,
and fractures developed as stresses were applied to the rocks
during orogenic and other geological processes.
Soil Mechanics / Geotechnical Engineering is concerned with
the engineering behaviours of earth materials (i.e. soils, and
weathered rock).
Difference of Rock Mechanics and
Geotechnical Engineering
11. 11
(a) Suitability To assess the general suitability of the site
(b) Design To enable an adequate and economic design.
(c) Construction
(i) To plan the best method of construction;
(ii) To foresee and provide against difficulties and
delays that may arise during construction; and
(iii) To explore sources of indigenous materials for
use in construction.
(d) Effect of Change
To determine the changes that may arise in the
ground and environmental conditions.
Objective of Ground Investigation (GI)
14. 14
Ground Subsidence
Collapsed area: 100m by 130m; settlement up to 15m
Elura Mine, NSW, Australia
Source: http://en.wikipedia.org/wiki/Image:Elura.png#file
19. Ground Condition Risks
19
Fookes’ (1997) study indicated:
• ~50% (confidence) of the anticipated
geological model from desk study.
• ~65% (confidence) of the geology should be
known if a walkover survey is added to the
desk study.
• 95% (confidence) if comprehensive GI works
to be done.
21. 21
US National Committee on Tunnelling
Technology (1984) suggested:
• 1.5 linear metre of borehole per route metre
tunnel alignment, and
• ~3% of cost of tunnelling civil works for ground
investigation.
23. 23
GI for Hard Rock Openings
Source: AGS (HK)
DH(I)
DH(V)
24. 24
Typical Tests Required to Interpret Design Parameters
In Situ Tests:
SPT, Water absorption test, Packer
test, Lugeon tests, Impression
packer/BH televiewer
Geophysical surveys: seismic,
resistivity, micro-gravity, magnetic,
cross-hole shear wave test
In situ modulus: High Pressure
Dilatometer or Goodman Jack, etc
In situ stress tests (e.g. Hydraulic
Fracturing Test, Flatjack, Overcoring
Test, Pressuremeters, High
pressure dilatometer)
Laboratory Tests:
Index tests, Triaxial shear strength
and Oedometer for overburden
Point load, UCS, Young's Modulus,
Poisson's ratio, Rock shear test on
joints, shear-box test for joint, saw
cuts for rock, Modulus of rupture
of rock, etc.
Testing for TBM/Machinery selection:-
Thin section petrography, Punch test,
Rock abrasively test, Brazilian test,
Machine Excavation Performance test,
Cuttability & Drillability Test
Field and Laboratory Testing
25. 25
Rock Tunnel/Cavern Design Parameters
Geological model (desk study, GI)
Groundwater level, permeability of soil/rock mass
(GI, field testing)
Insitu Stresses (field testing)
Rock Mass Quality (e.g. RMR, Q, GSI) (field mapping,
rock cores inspection)
Joints orientations, shear strength (c’ & f’), stiffness
(field mapping, lab testing, empirical methods)
Rock and Rock mass strength, modulus, shear strength
(c’ & f’), Poisson's Ratio (field and lab testing,
empirical methods)
33. 33
Insitu Stresses Field
Rock at depth is subjected to stresses resulting
from the weight of the overlying strata and
from locked in stresses of tectonic origin.
When an opening is excavated in this rock, the
stress field is locally disrupted and a new set
of stresses are induced in the rock surrounding
the opening. (Hoek 2007)
34. 34
Insitu Stresses Field – Vertical Stress
sv = g · z
Where:
sv is the vertical stress
g is the unit weight of the overlying rock and
z is the depth below surface
(After Brown and Hoek 1978)
35. 35
Insitu Stresses Field – Horizontal Stress
Normally, the ratio of the average horizontal stress (sh)
to the vertical stress (sv) is denoted by the letter k
such that:
sh = k · sv = k · g · z
k = sh / sv
(Hoek et al 2000)
(ksv)
44. 45
Field Testing and Measurements of
Insitu Stresses
Method :
Flat Jack
Hydraulic Fracturing Test including
hydraulic tests on pre-existing
fractures
Overcoring Test
CSIR / CSIRO cell
Borre probe cell
USBM
Sigra IST
48. 49
Suggested method for deformability determination
using a large flat jack technique
J. Loureiro-Pinto
International Journal of Rock Mechanics and Mining Sciences & Geomechanics
Abstracts, Volume 23, Issue 2, April 1986, Pages 133-140
55. 56
Assumptions/Considerations of HF (or HTPF)
sv = gravity body force of rock at depth
Principal stresses orientated at true vertical
and horizontal
Test at shallow ground (i.e. < 30m) may give a
questionable results
57. 58
Instrument
No of active
gauges
Measuring
depths
Continuous
logging
Borehole
requirements
CSIR Cell 12
Normally 10–50 m;
modified versions
up to 1000m
No
38mm pilot hole, usually
90mm drillhole. Modified
versions accept water
CSIRO Cell 9 / 12
Normally up to
30m
Yes, via cable
38mm pilot hole, usually
150mm drill hole.
Problems in water filled
holes
Borre probe
cell
9
Practiced to 620 m.
Tested for 1000m
Yes, built in
datalogger
36mm pilot hole, 76mm
drillhole.
Accepts water-filled holes
USBM
Normally 3;
modified
versions 4
Normally 10–50 m;
modified versions
up to 1000m
No
38mm pilot hole, usually
90mm drillhole.
Modified versions accept
water
Sigra IST
3, in two or
three levels
Used to 700 m.
Designed for
1500m
Yes, built in
datalogger
25mm pilot hole, 76mm
drillhole.
Accepts water-filled holes
List of Overcoring Testing Cell
69. 70
Hemispherical Projection Method (also called Stereo-
graphic Projection), there are two projection methods:
Use of Hemispherical Projection Method
Equal Area Projection
Reducing areal distortion and improving visual
estimates of clusters and variabilities.
Equal Angel Projection
When performing kinematic analysis, angular
relationships and shapes are preserved.
72. 73
Jointing data of
Lower Road Slopes
Jointing Data
Of Upper Road
Jointing data of
Lower Road Works
Legend
Drillholes
Cut slopes
Fill slopes
Disturbed terrain
Jointing data of
Lower Road Slopes
Rock Joint Analysis Example
89. 90
Hoek-Brown Empirical Failure Criterion
For highly fractured
rock, it reduces in
value of “s” (i.e. < 1)
and tends towards
zero as the strength is
reduced from peak to
residual.
92. 93
If the discontinuity is parallel or perpendicular to the applied
loading, it will have no effect on the sample strength. If the
discontinuity orientated at some angles, it will significantly
reduce the strength of the sample.
Strength of Rock with Single Joint
Intact Rock
MC Model
93. 94
Mohr's Circle - Possible modes of failure for rock
containing a single plane of weakness.
Circle A represents the case when the failure locus for the discontinuity is
just reached, i.e. for a discontinuity at the angle 2bw = 90 + fw.
Circle B – For a case when failure can occur along the discontinuity for a
range of angles, as indicated in the figure.
Circle C – For the case where the circle touches the intact rock failure locus,
i.e. where failure will occur in the intact rock if it has not already done so
along the discontinuity.
95. 96
Strength of Jointed Rock
Each discontinuity would weaken the sample (as discussed in previous slide),
but the angular position of the strength minima would not coincide. As a
result the rock is weakened in several different directions simultaneously.
Hence, heavy jointed material tends to become isotropic in strength, like a
granular soil (Hudson & Harrison 1997).
96. 97
Isotropic medium Anisotropic medium
Strength of Jointed Rockmass
In most of numerical model, the geomaterials (soil/rock) are considered
to be Continuous, Homogeneous, Isotropic and Liner-Elastic (CHILE).
However, in reality the geomaterials are Discontinuous, Inhomogeneous,
Anisotropic and Non-Elastic (DIANE).
[e.g. GSI=30 (or Q=0.1); RQD=25]
100. 101
Shear Strength of Rough Surfaces
fb is the basic friction angle of the surface and
i is the angle of the saw-tooth face.
101. 102
Barton (1990) equations:
Where:
JRC = joint wall roughness coefficient
JCS = joint wall compression strength
sn = normal stress of the block
fb = basic friction angle of rock joint
Barton’s Estimate of Rock Joint Shear Strength
103. 104
Joint Wall Compressive Strength, JCS
Estimate of joint wall
compressive strength (JCS)
from Schmidt hardness
(after Barton et. al., 1977 and 1985)
Bandis et al (1983) suggested:
F to SW: (sc / JCS) ~< 1.2
MW: 1.2 < (sc / JCS) ~< 2
W: (sc / JCS) > 2
104. 105
Joint Wall Stiffness (Barton 1972)
For a single joint set with an average spacing L, oriented
perpendicularly to the direction of loading, the joint normal
stffness (kn) is:
𝒌 𝒏 =
𝑬𝒊 𝑬 𝒎
𝑳 𝑬𝒊 − 𝑬 𝒎
where Em = rock mass modulus; Ei = intact rock modulus,
Gm = rock mass shear modulus; Gi = intact rock shear modulus,
L = mean joint spacing.
Joint shear stiffness (ks) is:
𝒌 𝒔 =
𝑮𝒊 𝑮 𝒎
𝑳 𝑮𝒊 − 𝑮 𝒎
𝐺 =
𝐸
2 1 + 𝑣
105. 106
Rockmass Permeability –
Water Ingress Assessment for
Underground Openings
Reference:
Kong, W.K. 2011. Water Ingress Assessment for Rock Tunnels: A Tool for
Risk Planning. Rock Mechanics and Rock Engineering, Volume 44, Number
6, pp. 755-765.
Open access to download:
http://link.springer.com/article/10.1007/s00603-011-0163-4?view=classic
107. 108
Rockmass Classification
Terzaghi's rockmass classification (Terzaghi, 1946)
Geomechanics Classification or the Rock Mass Rating
(RMR) system (Bieniawski, 1976)
Rock Tunnelling Quality Index, Q (Barton et al, 1974)
Geological strength Index (GSI) (Hoek ,1994)
108. 109
Terzaghi's rockmass classification (1/2)
Rock class Type of Rocks Definition
I Hard and intact
The rock is unweathered. It contains neither joints nor hair cracks. If
fractured, it breaks across intact rock. After excavation, the rock
may have some popping and spalling failures from roof. At high
stresses spontaneous and violent spalling of rock slabs may occur
from the side or the roof. The unconfined compressive strength is
equal to or more than 100 MPa.
II
Hard stratified and
schistose
The rock is hard and layered. The layers are usually widely separated.
The rock may or may not have planes of weakness. In such rocks,
spalling is quite common.
III
Massive,
moderately jointed
A jointed rock, the joints are widely spaced. The joints may or may
not be cemented. It may also contain hair cracks but the huge blocks
between the joints are intimately interlocked so that vertical walls
do not require lateral support. Spalling may occur.
IV
Moderately blocky
and seamy
Joints are less spaced. Blocks are about 1m in size. The rock may or
may not be hard. The joints may or may not be healed but the
interlocking is so intimate that no side pressure is exerted or
expected.
V
Very blocky and
seamy
Closely spaced joints. Block size is less than 1 m. It consists of almost
chemically intact rock fragments which are entirely separated from
each other and imperfectly interlocked. Some side pressure of low
magnitude is expected. Vertical walls may require supports.
109. 110
Terzaghi's rockmass classification (2/2)
Rock class Type of Rocks Definition
VI
Completely
crushed but
chemically intact
Comprises chemically intact rock having the character of
a crusher-run aggregate. There is no interlocking.
Considerable side pressure is expected on tunnel
supports. The block size could be few centimeters to 30
cm.
VII
Squeezing rock –
moderate depth
Squeezing is a mechanical process in which the rock
advances into the tunnel opening without perceptible
increase in volume. Moderate depth is a relative term
and could be from 150 to 1000 m.
VIII
Squeezing rock –
great depth
The depth may be more than 150 m. The maximum
recommended tunnel depth is 1000 m.
IX Swelling rock
Swelling is associated with volume change and is due to
chemical change of the rock, usually in presence of
moisture or water. Some shales absorb moisture from air
and swell. Rocks containing swelling minerals such as
montmorillonite, illite, kaolinite and others can swell and
exert heavy pressure on rock supports.
110. Terzaghi Rock Load
16 March 2011
111
Support pressure (pv) = g · Hp
where g is unit weight of rock
111. Terzaghi Rock Load
112
Comments on Terzaghi Rock Load
• Terzaghi’s method provides reasonable support
pressure for small tunnels (B < 6 m).
• It provides over-safe estimates for large tunnels and
caverns (Diam. 6 to 14 m) and
• The estimated support pressure values fall in a very
large range for squeezing and swelling ground
conditions for a meaningful application.
114. 115
RMR System [Bieniawski (1973, 1974, 1989)]
Six parameters are used to classify a rock mass using
the RMR system:
1) Uniaxial compressive strength of rock material
2) Rock Quality Designation (RQD)
3) Spacing of discontinuities
4) Condition of discontinuities
5) Groundwater conditions
6) Orientation of discontinuities.
RMR Rating = (1) + (2) + (3) + (4) + (5) + (6)
123. 124
RMR System - Guidelines for excavation and support
of 10 m span rock tunnels (After Bieniawski 1989)
124. RMR Support Pressure
125
Unal (1983), particularly applicable for flat roof coal mine
with span < 10m.
Goel and Jethwa (1991) short-term support pressure
for underground openings, but not for rock burst
condition.
where H = overburden or tunnel depth in meters (50–600 m)
127. 128
Q-System
SRF
J
J
J
J
RQD
Q w
a
r
n
where RQD is the Rock Quality Designation
Jn is the joint set number
Jr is the joint roughness number
Ja is the joint alteration number
Jw is the joint water reduction factor
SRF is the stress reduction factor
128. 129
The first quotient (RQD / Jn), representing the structure of
the rock mass, is a crude measure of the block size, with
the two extreme values (100/0.5 and 10/20).
The second quotient (Jr / Ja) represents the roughness and
frictional characteristics of the joint walls or filling
materials.
The third quotient (Jw / SRF) represents the active stress
in rock. Jw is a measure of water pressure with the effect
on the shear strength of joints due to a reduction in
effective normal stress. SRF is a measure of:
1) loosening load in the case of an excavation through
shear zones and clay bearing rock,
2) rock stress in competent rock, and
3) squeezing loads in plastic incompetent rocks.
Q-System
129. 130
‘Q’ value can be ranging from 0.001 to 1000
corresponding to extremely poor to excellent
rock conditions
Q-System
130. 131
Barton et al (1974) defined an additional parameter which
they called the Equivalent Dimension, De:
Q-System
Excavation Support Ratio (ESR) determined by:
131. 132
Estimate of Q Support Pressure
Barton et al. 1974 recommended:
𝐏 𝐯 =
𝟐𝟎𝟎 ∙ 𝐐−𝟏/𝟑
𝐉 𝐫
𝐏 𝐯 =
𝟐𝟎𝟎 ∙ 𝑱 𝒏
𝟏/𝟐 ∙ 𝐐−𝟏/𝟑
𝟑 ∙ 𝐉 𝐫
(kPa), if no. of joint sets ≥ 3
(kPa), if no. of joint sets ≤ 3
Bolt length (m). B is span or height
of opening whichever is larger.2+
132. 133
Range of Q
Wall Roof
Factored
Qwall
Temporary
Qt
wall
Temporary
Qt
roof
Q > 10 5.0 x Q 5 x 5 x Q 5.0 x Q
0.1 < Q < 10 2.5 x Q 5 x 2.5 x Q 5.0 x Q
Q < 0.1 1.0 x Q 5 x 1.0 x Q 5.0 x Q
Q-value Adjustment for Tunnel Wall and
Temporary Conditions of Opening
For temporary case (< 1 year), ESRtemp = ESR x 1.5
133. 134
Q Design Chart (NGI, 2015)
Bolt spacing based on f20mm dia.
but design working load ???
100kN?
141. 142
Correlation between Systems
RMR = 9 In Q + 44 (Bieniawski, 1989)
RMR = 15 log Q + 50 (Barton 1995)
GSI = RMR89 – 5 (Hoek & Brown 1997)
Warning: The Q-system and the RMR system include
different parameters and therefore cannot be strictly
correlated. Palmström & Stille (2010), the relationship
has an inaccuracy of ± 50% or more.
144. 145
Leontovich (1959) gives solution for arches with raise-to-span
ratio (r/Span) ranging from 0 to 0.6 for which the
recommended assumptions for loading of such arches are
believed to be safe:
For low rise arches (r/Span) = 0.2 or less, a uniform load
may be assumed.
For higher rise arches (r/Span) > 0.2, a dead load consisting
of uniform plus complementary parabolic loading (similar
to Terzaghi’s rock load) may be assumed..
Principle of Roof Arch Depth
145. Generic Structural Arch Beam Formula
146
Symmetrical Three-Hinged Arches of any Depth
(Milkhelson 2004)
151. Terzaghi Rock Load
152
Comments on Terzaghi Rock Load
• Terzaghi’s method provides reasonable support
pressure for small tunnels (B < 6 m).
• It provides over-safe estimates for large tunnels
and caverns (Diam. 6 to 14 m) and
• The estimated support pressure values fall in a very
large range for squeezing and swelling ground
conditions for a meaningful application.
152. Comments on Terzaghi Rock Load (cont’)
153
Barton et al. (1974) and Verman (1993) suggested
that the support pressure is independent of opening
width in rock tunnels. Goel et al. (1996) also found
that there is a negligible effect of tunnel size on
support pressure in non-squeezing ground
conditions, but the tunnel size could have
considerable influence on the support pressure in
squeezing ground condition.
153. RMR Support Pressure
154
Unal (1983) for flat roof coal mine
Goel and Jethwa (1991) short-term support pressure
for underground openings, but not for rock burst
condition
where H = overburden or tunnel depth in meters (50–600 m)
154. Q Support Pressure
155
Barton et al. 1974 recommended:
𝐏 𝐯 =
𝟐𝟎𝟎 ∙ 𝐐−𝟏/𝟑
𝐉 𝐫
𝐏 𝐯 =
𝟐𝟎𝟎 ∙ 𝐉 𝐧
𝟏/𝟐
∙ 𝐐−𝟏/𝟑
𝟑 ∙ 𝐉 𝐫
(kPa), if no. of joint sets ≥ 3
(kPa), if no. of joint sets ≤ 3
158. 159
Q-system Design Approach:
1) Determine an opening size (height, span)
2) Determine Excavation Support Ratio (ESR)
Tunnel/Cavern Support Design in Rock
159. 160
3) Calculate roof and wall supporting stress for
different Q-values
Q-system Design Approach (cont’):
𝐏 𝐯 =
𝟐𝟎𝟎 ∙ 𝐐−𝟏/𝟑
𝐉 𝐫
𝐏 𝐯 =
𝟐𝟎𝟎 ∙ 𝐉 𝐧
𝟏/𝟐
∙ 𝐐−𝟏/𝟑
𝟑 ∙ 𝐉 𝐫
(kPa), if no. of joint sets ≥ 3
(kPa), if no. of joint sets ≤ 3
Range of Q
Wall Roof
Factored
Qwall
Temporary
Qt
wall
Temporary
Qt
roof
Q > 10 5.0 x Q 5 x 5 x Q 5.0 x Q
0.1 < Q < 10 2.5 x Q 5 x 2.5 x Q 5.0 x Q
Q < 0.1 1.0 x Q 5 x 1.0 x Q 5.0 x Q
160. 161
4) Determine bolt length, bolt force and bolt spacing
Q-system Design Approach (cont’):
Bolt length (m). B is span or height of opening
whichever is larger.
For temporary case (< 1yr), ESRtemp = ESR x 1.5
Design Working Load of Bolt ≤ 0.5 x characteristic yield
strength of bolt (e.g. BS 8081)
Bolt Spacing = (Support pressure / Working Load of Bolt)0.5
2+
163. The concept is based on improving the strength of the
rockmass at the tunnel walls by application of confining
pressure via the bolts.
What is Rock Reinforcement for
Underground Opening in Hard Rock
164(Bischoff and Smart, 1975)
164. How does it work of Rock Reinforcement?
165
Photoelastic stress pattern of bolting
Lang’s (1961) findings:
165. 166
How does it work of Rock Reinforcement?
(excerpted from Hoek, 2007)
166. How does it work of Rock Reinforcement?
167
Theoretical zone of compression by bolting (Hoek, 2007)
where, L ≈ 2 to 3 S; and S < 3a
167. Concept of Reinforcement of Rock Arch
168
Where: σc is the unconfined compressive strength of rockmass
σt is the tensile strength of rockmass (by consideration of MC criterion)
Fb is provided bolt force
(in half span tunnel, kN)
(Bischoff and Smart, 1975)
168. SSR – The shotcrete liner is designed as a structural
liner to support a failure wedge occurred between
bolts. Detailed study on the SSR and structural
shotcrete liner design has been carried out by number
of researchers, they are:
• Fernandez-Delgado et al (1981)
• Holmgren (1987)
• Vandewalle (1992)
• Barrett & McCreath (1995)
• Morton et al (2009)
• Uotinen (2011) compatible with Eurocodes
• Kong & Garshol (2015)
Shotcrete-Rock-Reinforcement (SRR)
169
170. Failure modes of SSR
171
Adhesion Failure
Shear FailurePunching Failure
Flexure Failure
[modified from Barrett and McCreath (1995)]
unstable wedge between bolts, 45° projected from the base plate of the rock bolt (for critical case)
171. 172
Determining Adhesive Failure of SSR
Based on Barrett and McCreath (1995), and Uotinen (2011)
Barrett and McCreath (1995) carried out back-calculations to reveal
that the required adhesion strength for high grade shotcrete in hard
rock was typically 0.5 MPa, and a minimum of 30 mm conservative
bond width may be used in the design. If the adhesion strength is
unknown, 0.4 MPa may be used for a conservative case.
The Rad should be designed strong enough to retain a potential rock
wedge forming in between rock bolts.
Where: fak is the adhesion strength (bond strength in MPa)
S is the perimeter of the load to be supported (i.e. bolt spacing in metres)
b is width (in metres) of the adhesion area (if unknown, 30 mm may be used)
γc is the partial safety factor for concrete (BS EN 1992-1-1:2004 s. 2.4.2.4)
172. 173
Bending Capacity of Shotcrete Liner
Where: σflex is the pure bending tensile strength
[see Eq. (3.23) of BS EN 1992-1-1:2004]
t is the shotcrete thickness (m)
Designed Moment of Shotcrete Liner
Where: w is the contributed load of failure wedge (kN)
S is the bolt spacing (m)
c is width of the faceplates (m)
Cflex > Mo
Determining Flexure Failure of SSR
Based on Barrett and McCreath (1995), and Uotinen (2011)
173. 174
Shear Capacity of Shotcrete Liner
Designed Shear Failure of Shotcrete Liner
Where: fctm is the shear (or tensile) strength (in MPa) of shotcrete grade
S is the perimeter of the load to be supported (i.e. bolt spacing in metre)
t is the thickness of the shotcrete layer (m)
γc is the partial safety factor for concrete (BS EN 1992-1-1:2004 s. 2.4.2.4)
Where: w is the contributed load of failure wedge (kN)
S is the bolt spacing (m)
Rvd > Rsd
Determining Shear Failure of SSR
Based on Barrett and McCreath (1995), and Uotinen (2011)
174. 175
Punching Shear Resistance of Unreinforced Shotcrete
Designed Punching Shear acting on Shotcrete Liner
It may follow the requirements as stipulated in Section 6.4.4(1) of
BS EN 1992-1-1:2004, to determine punching shear resistance, VRd,c
Where: w is the contributed load of failure wedge (kN)
S is the bolt spacing (m)
c is width of the faceplates (m)
V = w · (S² – c²)
VRd,c > V
Determining Punching Shear Failure of SSR
Based on Barrett and McCreath (1995), and Uotinen (2011)
178. 179
What’s wrong of the models?
After rockbolts were installed, no joint networks should be
added in the “Zone of Compression” to assess structural
response of shotcrete lining. This zone is treated to become
a continuous media and isotropic in strength.
Whatever it is PHASE2
or UDEC model, the
failure wedge size is
not true (Kong et al
2016), and hence the
shear force given by
the model is a
reference value only
(i.e. not a true value).
179. 180
Comparison with
Q-system and Rock Reinforcement
RR
(Lang, 1961; Bischoff & Smart, 1975)
Q-System
Bolt length
Depends on:
• rock strength,
• block size of rockmass,
• bolt spacing
• Bolt force
Depends on:
• Q value
• Opening size
• Excavation Support
Ration(ESR)
L = 2 + (0.15B/ESR)
Stablisation of
individual failure
wedge
Not considered Not considered
Rockmass strength
< 25 MPa
Application Questionable Applicable Questionable
Shotcrete liner
Structural liner against failure
wedge occurred between bolts
Prescribed thickness based
on past experience.
183. 184
Pillar Strength Estimation
𝝈 𝒑 = 𝑲
𝑾 𝜶
𝑯 𝜷
Pillar Strength (Salamon and Munro, 1967):
where sp (MPa) is the pillar strength,
K (MPa) is the strength of a unit volume
of rock
W and H are the pillar width and height
in metres
For square pillar
𝑾 𝒆 = 𝟒
𝑨 𝒑
𝑹 𝒑
where We = effective pillar width (m)
Ap = cross-section area of pillar (m²)
Rp = pillar circuference (m)
Wagner (1980) and Stacey & Page (1986) proposed:
For rectangular pillar
185. 186
Pillar Stress Determination
𝝈 𝒔 =
(𝒂 + 𝒄)∙(𝒃 + 𝒄)∙𝑯∙𝜸
(𝒂∙𝒃)
Pillar Stress:
Where:
a and b is the pillar
dimension
c is extraction width
H is the total depth of rock
above pillar
g is the unit weight of rock
186. 187
Pillar Stability
Factor of Safety (FoS) =
𝜎 𝑝
𝜎𝑠
Numerical modelling (e.g. PHASE2) is able to give a FoS of
Pillar Stability.
Where: sp is pillar strength
ss is pillar stress
Suggested FOS for Pillar Stability:
1.6 Salamon and Munro (1967) (coal mines pillar study)
>1.5 Hoek & Brown (1980), after Salamon and Munro (1967)
1.4 Lunder and Pakalnis (1997), and width to height ratios of
up to 1.5
1.4 Martin & Maybee (2000)
Pillar Stability:
189. 190
Crown Pillar Safe Span Estimation
S =
2𝑅𝑑
𝛾𝐹
Considers as Fixed End Roof Beams (Adler and Sun, 1968):
Where:
S – Safe Roof Span (m)
R – Modulus of Rupture of Rock (MPa)
(rock testing refers to ASTM C99/C99M-2015)
d – Thickness of Roof Beam
F – Factor of Safety (from 4 to 8)
Numerical modelling to verify overall stability of caverns
including crown pillar is required under different load
cases and the influences of insitu stresses.
190. 191
Design Chart of Safe Span vs Roof Beam
Thickness (Adler and Sun, 1968)
R=6R=4 R=8
R=10
R=2
(e.g. Granite)
