- Email: [email protected]
Optimising blast pulls and controlling blast-induced excavation damage zone in tunnelling through varied rock classes
Tunnelling and Underground Space Technology 85 (2019) 307–318
Contents lists available at ScienceDirect
Tunnelling and Underground Space Technology journal homepage: www.elsevier.com/locate/tust
Optimising blast pulls and controlling blast-induced excavation damage zone in tunnelling through varied rock classes Amiri Hamis Saluma,
⁎,1
a b
T
, V.M.S.R. Murthyb
Former Postgraduate student at Indian Institute of Technology (Indian School of Mines), Dhanbad, India Professor & Head of Department, Department of Mining Engineering, Indian Institute of Technology (Indian School of Mines) Dhanbad, Jharkhand 826004, India
ARTICLE INFO
ABSTRACT
Keywords: Rock mass characterization Round length Excavation sequence Overbreak Explosive charge quantity
Tunnelling is the sole means of providing access for transportation, water conveyance in hydropower, mining of minerals, etc. Hydropower alone contributes to about 19% of the commercial energy needs of India and construction of hydroelectric power projects to meet these needs involves huge amounts of tunnelling through varied rock mass conditions. This necessitates identifying methods of tunnelling that are cost-effective, suited to varied geology and yet rapid to commission the projects in place of continuous excavation systems like TBM and Roadheader which are geology sensitive, costly and completely imported thus constraining their smooth adaptation. Mechanized drilling and advanced blasting techniques are the most often used methods of excavation of tunnels in India and in other countries for rapid and cost effective tunnelling, which depends, to a large extent, on the blast performance. Among the outcomes of any blasting operation, pull or advance achieved per blast and blast-induced excavation damage are of major concerns. It is essential to limit the blast-induced damage so as to control time and cost overruns in an underground project more so in varied geology. This paper discusses determination of optimum round lengths of excavation in varied rock classes as well as controlling overbreak in tunnelling. The sequence of excavation, requirements for both top heading and bench are also addressed. It has been observed from the past experiences that round lengths up to 5 m are practiced in rock class I and as the class improves to VII, the pull attempted reduces to about 1 m. Since charging of perimeter holes contributes to overbreak in underground excavations, a thorough analysis of the design of blasting pattern and scheme of charging for minimizing overbreak has been suggested. Characterization of ground through seismic imaging coupled with ground vibration monitoring has been suggested to control blast – induced rock damage and also arrive at optimum charges. As evident in this paper, characterization of the ground is the most important step towards rapid tunnelling.
1. Introduction Drilling and blasting method is generally inevitable for rock excavation activities in mining, quarrying and civil construction works. Use of explosive energy probably is the most widely used means of crushing rock, as well as the most cost-effective rock excavation method (Yilmaz and Unlu, 2014). Compatibility and feasibility to any sudden required alteration in dimension of excavation profile and/or geological constraints also adds to the popularity and suitability of this method over any other methods of excavation such as TBM’s, Road headers and Impact hammers (Mandal and Singh, 2009). However, performance of drilling and blasting method is mainly influenced by a number of factors that can be categorized as: Rock mass features, explosives characteristics and their distribution, blast design
and execution. For a given tunnel alignment, rock mass features cannot be changed but their knowledge facilitates the judicious selection of the drilling systems and patterns, explosive characteristics and the blast design parameters to obtain optimum pull and reduced excavation damage (Singh and Xavier, 2005). The tunnel blast performance is generally measured in terms of one or more than one of the following blast parameters: 1. Pull (face advance/depth of round), expressed in percent, 2. Specific charge (kg of explosive/m3 or t of yield), 3. Specific drilling (m of drilling/m3 or t of yield), or Detonator or hole factor (number of holes/m3 or t of yield), and 4. Blast-induced rock mass damage or overbreak or underbreak (Chakraborty et al., 2004).
Corresponding author. E-mail addresses: [email protected] (A.H. Salum), [email protected] (V.M.S.R. Murthy). 1 Currently at: Tunnel Engineer, Unitec Civil Consultant Ltd, P.O.Box 32507, Dar-es-Salaam, Tanzania. ⁎
https://doi.org/10.1016/j.tust.2018.11.029 Received 20 May 2016; Received in revised form 28 July 2017; Accepted 24 November 2018 Available online 02 January 2019 0886-7798/ © 2018 Elsevier Ltd. All rights reserved.
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
2. Rock mass characterization for tunnel blasting The main focus in tunnelling projects is to achieve longer pulls (up to 6 m) considering the large tunnel size available. Currently, pull achieved in different blasting practices ranges from 1.5 m to 4 m and is shown in Fig. 1. In order to achieve longer pull with minimum rock damage, there is a need to go for thorough rock mass characterization so as to develop a proper blast design in terms of blast geometry, pull length, powder factor and choosing the right type of explosive for controlling the blastinduced excavation damage zone (EDZ). The interaction of parameters is shown in Fig. 2. Rock mass is a heterogeneous material, a fact rarely considered and properly represented in blast design. In reality rock mass features have a controlling influence on the outcome of a blast (Singh and Xavier, 2005). The rock mass parameters which influence blast results fall into two groups. The first group is the intact rock properties and the second group is the discontinuity structure (Ryu et al., 2006). For the purpose of blast design, it is very important to consider the physico-mechanical including dynamic properties of rock due to the fact that under dynamic loading, the strength of rock is reported to vary with the rate of loading or strain rate. The dynamic properties of rocks to be considered during design are: Young’s modulus (E), Poisson’s ratio (ν), compressive strength (σc), dynamic tensile strength (σtd), P-wave velocity (Cp), and S-wave velocity (Cs). Other static properties include density of rock (ρr) and specific gravity (Y). Rock mass features like orientation of discontinuities, aperture of discontinuities, frequency of discontinuities and filling in the joints contribute significantly in producing damage to the perimeter of an excavation. Rock Quality Designation (RQD) also gives fair indication of the quality of rock mass (Singh and Xavier, 2005).
Fig. 1. Pull achieved in various drilling and blasting operations in tunnelling and mining.
2.1. Rock mass classification vis-à-vis tunnelling There are different classification systems used to describe the rock mass quality. Among these classification systems, Rock Mass Rating (RMR), Qsystem and New Austrian Tunnelling Methods (NATM) are widely used for excavation and support design in tunnelling. However, in this study the following classes and ranges as shown in Table 1 are adopted. In the table, ten classes of rocks based on NATM classification and their approximate range in Q-system and RMR are described. Behaviour of each class is also explained to give complete understanding of the rock class. Method of rock excavation, especially, by drilling and blasting varies depending on the quality of rock and the likelihood of rock damage. Typical drilling and blasting schemes for tunnelling through varied classes are shown in Fig. 3. Normal blasting can be used in class I and II with fairly good results. But rocks belonging to Class III through V require cautious blasting. Class VI and VII require pre support like forepoling, baby arch and low strength explosives. Excavation of rocks belonging to Class VIII and below, make use of very light blasting or mechanical excavation (Mathis and Page, 1995).
Fig. 2. Blasting system in tunnelling and interaction of parameters.
Two main objectives of any tunnel blasting operation are to get higher drivage rates and minimize damage to the surrounding rock mass. Attempts to get higher pull sometimes lead to roof rock damage, more so in case of weak and jointed rocks. In order to optimize blastinduced rock damage, rock mass characterization and assessment of the extent of rock damage are the pre requisites. Most of the existing criteria relate damage to ground vibrations resulting from dynamic stresses induced by blasting practices (Murthy and Dey, 2004). Thus, it is very important to characterize the rock mass for blasting purposes in order to achieve optimum results. This paper presents a methodology of characterizing the rock mass, various theoretical approaches to assist in achieving higher pull and minimize damage in tunnelling.
Table 1 Approximate relationship between NATM, Q-system and RMR system (Karahan, 2010). Class
NATM system term
Q-system
RMR
Remarks
I II III IV V VI VII VIII IX X
Stable Slightly Overbreaking Friable Very Friable Rolling Rock Bursting Squeezing Heavily Squeezing Flowing Swelling
> 70 10–70 4–10 1–4 0.11–1 0.03–0.11 0.015–0.03 0.008–0.015 0.002–0.008 < 0.002
> 80 65–80 58–65 47–58 29–47 20–29 15–20 10–15 5–10 <5
The rock mass is permanently stable without support A slight tendency of shallow overbreaks in the tunnel roof and in the upper portions of the sidewalls Overbreaks and loosening of the rock strata in tunnel roof and upper sidewalls if no support is installed in time Stand-up time and unsupported span are short Failure mechanisms such as spalling, buckling, shearing and rupture of the rock structure Rapid and significant movement of the rock mass into the cavity Limit the unsupported spans at arch and face Prior installation of forepoling or forepiling and shotcrete sealing of faces. The low cohesion requires a number of subdivisions
308
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
(a) Class I – II
(b) Class III – V
(d) Class VIII - X
(c ) Class VI – VII
Fig. 3. Illustration of different drilling and blasting techniques.
induced stresses, regional geology, structural geology and weak zone characteristics (Yu and Chern, 2007). It is possible to excavate full face for tunnels in Class I and II provided that the face area is within the reach of the drilling jumbo. But due to large cross section of some tunnels like motorway tunnels, staged excavation (Top heading and bench) is preferred in practice. However, in rock classes V and below multistage (Side drifting) is required for the stability of tunnels (Ulukan et al., 1998). Also an invert is usually provided for stability purpose (Chen and Liu, 2007). Yu and Chern (2007) have proposed a diagram that can be used as a basis for selection of excavation method (Fig. 4). Thus, the excavation method can be determined according to the given span and ratio of Uniaxial Compressive Strength (UCS) to vertical in situ stress. This concept was used in the design of excavation sequence in Niayesh Tunnel Project in Iran. This was the biggest tunnelling project in urban area in Middle East for its length, cross section area and step of the route. The tunnel project is mouth shaped, and is 13 m high and 18 m wide. Using Yu and Chern graph, it was revealed that Central diaphragm and side drifting methods are appropriate excavation patterns (Fig. 5). However, side drifting was preferred due to low settlement as observed from numerical modelling results of the two (Bolghonabai et al., 2015).
Fig. 4. Empirical determination of excavation sequences based on span size, unconfined compressive strength (UCS) and vertical stress on tunnel (Yu and Chern, 2007).
2.2. Determination of excavation method and sequences Generally, there is no simple rule to facilitate decision making about optimal selection of excavation method. However, some influencing factors on selection of excavation method can be mentioned considering rock mass properties including intact rock and joint (discontinuity) characteristics, shape and size of tunnel section, underground hydrology, in situ and
3. Determination of advance per round in tunnelling Traditionally, the advance per round is designed based on tunnel 309
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
(b) Central diaphragm (CD)
(a) Side drift
Fig. 5. Excavation stages selection for Niayesh Tunnel (Bolghonabai et al., 2015).
(a) Wedge cut in a small tunnel
(b) Parallel cut in small tunnel
Fig. 6. Variation in blasting parameters in different cuts practiced in tunnelling.
hydraulic radius in wedge cut and relief hole diameter in parallel cut. A wedge cut is more productive than parallel cut in small tunnels (Area < 10 m2) where the desired advance per round does not exceed 0.7 of the tunnel width. Here an advance rate of 0.4–0.7 times hydraulic radius can be obtained. (Chakraborty et al., 1998) On the other hand, the maximum advance per round in a parallel cut is influenced by the size of the relief holes but not the size of the opening. It is given by the following formula:
H = 0.15 + 34.1Ø
39.4Ø2
[Ø
0]
investigate the variation of pull with rock quality for eight classes (Table 2). Decision with regards to the round length must be verified with the RQD values as well as the sonic velocity of rocks. This is to ensure proper breaking of rock, by studying the impedance characteristics of the rocks so as to idealize the whole concept of energy transfer in the rock mass. Sjogren et al. (1979) studied RQD values based on sonic velocity data from various rock types and different geographical locations to analyze some of the relationships between seismic variables and their correlations with various physical properties of rock masses. Variation in cracks frequency and sonic velocity with RQD can be obtained directly from Fig. 7. RQD is an indicator of the quality of the rock mass. RQD of < 70% indicates that the rock mass will be more susceptible to blast damage and the RQD values < 50% would require close spacing, light loading and relief holes to produce acceptable results of overbreak (Singh and Xavier, 2005).
(1)
where H is the maximum advance per round, m, and Ø is the relief hole diameter, m. The parallel cut is less productive than the wedge cut in small size tunnels where the ratio of cut to opening area tends to one (Fig. 6). Higher charges and more number of blast holes are required in parallel cuts compared to a wedge cut. However, a parallel cut is more feasible in medium (Area = 10–35 m2) to large tunnels (Area > 35 m2) where the need of the higher quantity of charge and number of blast holes in the cut area can be neutralized by a higher ratio of the area between the tunnel opening and the cut section (Fig. 6). In parallel cut, usually a minimum of 25% of the volume of the initial cut should be removed by drilling relief holes (Adhikari et al., 1999). Pull designed in this way, does not take into account the rock quality which should be given due concern while designing tunnel blast. Data regarding pull were collected from about 40 case studies to
4. Blast vibrations and rock mass damage. Blasting damage in a rock mass is a result of the borehole strain developed by an explosive in a blast hole. According to Hooke’s law, the axial stress is related to the axial strain. This relationship can be expressed as:
= /E
(2)
where E is the elastic modulus of the material and σ and ε are the axial stress and strain respectively. 310
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Table 2 Blast round lengths in varied rock classes. S. No.
1 2 3 4 5
Excavation classes NATM CLASS Q system RMR Round length top heading (m) Round length of bench (m) Round length top heading (m) Round length top heading (m) Round length of Bench height (m) Round length top heading (m) Round length of Bench height (m) Round length for top heading (m) Round length for top heading (m)
Ref. I > 70 > 80 4.2 – <4
5 3.5
II 10–70 65–80 3 4.5 <3
4, 4.3, 3.2, 3 2.3, 2, 1.5
III 4.0–10 58–65 2.85 3.8 < 2.5 2–3 4 2–3 4 3 1.5, 2.5
IV 1.0–4.0 47–58 2.3 3.25 1.5–2 1.5–2.5 3.5 1.5–2 3–3.5 3 1.5
4.1. Rock damage assessment from peak particle velocities
(3)
Assuming brittle failure mode of the rock, a threshold of critical peak particle velocity PPVcrit which can be withstood by the rock before tensile failure can be computed from Eq. (4).
PPVcrit = (
t
VII 0.015–0.03 15–20 1.1 1.1 –
VIII 0.008–0.015 10–15
Ulukan et al. (1998) Chen and Liu (2007) Kockar and Akgun (2003)
0.75–1.25 <2
Kockar and Akgun (2003) From various projects
PPV (mm/s)
Tensile stress (MPa)
Strain energy (J/ Kg)
Typical effect in hard Scandinavian bedrock
700 1000 2500 5000 12,000
8.7 12.5 31.2 62.4 187
0.25 0.5 3.1 12.5 112.5
Incipient swelling Incipient damage Fragmentation Good fragmentation Crushing
(4)
Cp)/ E
Several researchers investigated the ground vibration levels during blasting in tunnels, quarrying and mines and suggested different threshold levels (PPV) for rock mass damage. The rock damage criteria suggested are presented in Table 4. Li and Huang (Wu and Hao, 2006) discussed PPV damage criterion for rock tunnels with the following definition: slight damage (initial cracking); medium damage (partial collapse); serious damage (largearea tunnel collapse). The respective PPV for the various types of rock mass are shown in Table 5. It may be observed that the average tensile strength values appear to relate well with the allowable threshold levels and this may be due to the breakage of rock being effected in reflection mode taking tensile strength as the key weakness. The relation is shown in Fig. 8. Thus, for a given tensile strength of rock the damage threshold levels (PPV) can be fixed and accordingly the charges can be controlled. On the other hand Singh (2012) suggested critical PPV values for different rock masses based on Compressive strength properties as given in Table 6. Either of tensile or compressive strength can be used to predict
where σt is the tensile strength. In addition, the vibration level above which some damage may be expected PPVmax can be estimated from Eq. (5).
PPVmax = 1.2[ t /(Cp
VI 0.03–0.11 20–29 1.15 1.5 0.8–1.2
Table 3 Effect of PPV levels in Scandinavian bedrock (Wu and Hao, 2006).
The peak strain in the rock mass can also be related to the PPV and compressional elastic wave velocity Cp as given by:
= PPV / Cp
V 0.11–1.0 29–47 – – 1–1.5 1.5–2 2.5 1.25–1.5 <3 2.5 1.7, 2, 2.3
r )]
(5)
where ρr is the density of rock in kg/m3 (Yilmaz and Unlu, 2014). Eqs. (3) and (4) can be used for determination of critical strain in rock mass of certain P-wave velocity, which is a characteristic of that medium. Knowledge of this critical strain will be used as a criterion for deciding the optimum or maximum pull for rock mass having similar properties. In addition to that, Eq. (4) is very important in determination of damage depth hence deciding the support requirement. Holmberg and Persson (1979) determined the limiting PPV based on new crack formation in the surrounding rock mass and they found that the critical peak particle velocity for new crack formation is between 700 and 1000 mm/s for Swedish hard igneous rocks (Holmberg and Persson, 1979; Wu and Hao, 2006). Table 3 shows the observed typical effects that would occur when the peak particle velocities vary in Scandinavian bedrock.
Fig. 7. Variation of cracks frequency and P-wave velocity with RQD. 311
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Table 4 Suggested damage threshold for rock mass damage (Murthy and Dey, 2003). Model (Year)
Damage criteria
Edwards and Northwood (1960), Langefors et al. (1973), Nicholls et al. (1971) Bauer and Calder (1970)
Langefors and Kihlstrom (1973) Holmberg and Persson (1979) Oriard (1982) Rustan (1985) Meyer and Dunn (1995) Bogdanhoff (1996) Murthy and Dey (2002) Dey (2004)
PPV (mm/s)
Damage description
< 50 < 254 254–635 635–2540 > 2540 305 610 700–1000 > 635 300–900 1000–3000 300 600 2000–2500 2050 700–1300
Low probability of structural damage to residential buildings No fracturing of intact rock Minor tensile slabbing Strong tensile slabbing and radial cracking Break up of rock mass Results in fall of rocks Results in formation of new cracks Rock mass damage Rock mass damage Smooth blasting Rock damage threshold Minor damage Damage threshold Damage threshold Threshold for overbreak in compact basalt Threshold for overbreak
Tunnelling Mining
Tunnelling Tunnelling Tunnelling Tunnelling Mining Tunnelling Tunnelling Metal mines
4.2. Blast-induced damage assessment using seismic imaging and PPV measurement
Table 5 Rock tunnel damage criteria (modified after Li and Huang (1994)). Average
Application
Peak particle velocity (mm/s)
Compressive strength (MPa)
Tensile strength (MPa)
No damage
Slight damage
Medium damage
Serious damage
92.5 145 190 70 130
2.75 4.25 5.4 2.1 3.95
270 310 360 290 350
540 620 720 580 700
820 960 1110 900 1070
1530 1780 2090 1670 1990
Blast-induced rock damage has been related with peak particle velocity (PPV) by many researchers and a mathematical formula has been proposed for assessing the extent of damage zone by extending the formula proposed by Holmberg and Persson (1979) for the near-field vibration approximation (Murthy and Dey, 2003). This model gives damage envelope when plotted in space (x, y). To test the acceptability of the proposed model, a crater blast experiment has been carried out in the bench of a surface mine with 1 m long and 32 mm dia drill holes, at Nuasahi chromite surface mine in Orissa. The holes have been blasted with 250 g (0.4 m) explosive. Vibration monitoring has been carried out close to blast site to establish the vibration predictor. Seismic imaging has been carried out at the blast site before and after the blast. Analysis of the seismic images (pre and post blast) has been done to decipher the
critical PPV with satisfactory results, but authors of this paper would prefer using tensile strength as the same can be used in Tezuka’s equation to calculate PPV and strain.
Fig. 8. Effect of rock properties on threshold levels of damage (modified after Li and Huang, 2008). 312
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Trial blasts were conducted and vibration monitoring was done for establishing the ground vibration attenuation equation. For both the rounds of blast conducted, the PPV was measured for varying maximum charge per delay (W) and radial distance (R). Table 8 presents the details of data generated. For minimizing the blast-induced vibration and related rock damage, the blasts were executed in two rounds. In Round-I, holes up to spring level were blasted and the remaining holes were blasted in Round-II, after necessary dressing (Fig. 10).
Table 6 Critical velocities for different types of rocks. Rock type
Hard Medium soft
Properties Specific gravity
Uniaxial compressive strength (MPa)
> 2.7 > 2.5 > 2.3
> 240 100–150 < 50
Critical PPV (mm/ s)
1200–2000 700–1000 < 400
5.2. Determination of PPV predictor equations
extent of rock damage. Using the proposed model, the damage envelope has been computed. Comparing the damage envelopes obtained from the proposed model and seismic imaging, it has been found that they are in close agreement. Thus, it can be inferred that the suggested PPV based model holds good to decipher the damage zone. Apart from this, it has been found that the damage zones predicted from seismic images were 2–30 times larger than the physically measured crater volumes. This probably includes the cracked as well as incipient cracks zone. Seismic imaging technique has, thus, been found suitable for determining the damage extent with considerable accuracy. The pre and post blast Pwave velocity values were obtained and are shown in Fig. 9(a) and (b) respectively. It may be observed that this method can help assess the damage zone surrounding the blast hole there by optimizing the charges that can be exploded to control the overbreak.
Using data presented in Table 10, ground vibration predictor equations were established between PPV and scaled distance (SD). For Round-I, the scaled distance was given by the formula,
SD =
3
R W
(6)
Therefore, the PPV predictor equation was given as:
V = 514.98 ×
3
R W
1.2
,
Correlation coefficient r = 0.87
(7)
where V = PPV (mm/s) R = Distance (m) W = Maximum charge/delay (kg) In case of Round-II, the scaled distance was modified by introducing the stiffness factor, defined as the ratio of burden to hole depth, whose value is 0.6. Hence the best fit ground vibration equation obtained is:
5. Field investigations, data collection and analysis Blast design and ground vibration monitoring were carried out in Koyna Hydel Project by the authors and the same has been described below.
V = 93.95 ×
5.1. Blast design and vibration data monitoring
R 1 W 3 + SF
1.24
,
correlation coefficient R = 0.80
(8)
Critical velocity is given by Vcr = = 0.635 m/s. The value obtained is slightly less than that proposed by Singh and Holmberg and Persson (0.7 m/s). The damage distance obtained is 0.25 m as shown in Fig. 11. t × VP E
A pull of 1.5 m was chosen for considering cycle optimization and other factors such as lack of advanced initiation systems. The intact properties of rock are presented in Table 7.
a) Pre-blast seismic image
b) Post blast seismic image
Fig. 9. P-wave velocity profile surrounding a blast hole for blast-induced damage assessment (Dey and Murthy, 2011). 313
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
5.3. Overbreak control in tunnelling
Table 7 Important rock properties of lake tap horizontal tunnel, Koyna Project, India (Murthy and Dey, 2003). Rock characteristics
Value
Value taken in this study
Q-value P-wave velocity (m/s) UCS (MPa) Density (g/cm3) Tensile strength (MPa) Young modulus (MPa)
15–21 1970–5200 13–95 2.16–2.97 1.2–8.1 33,000–66,330
18 5200 95 2.97 8.1 66,330
Overbreak may be defined as the percentage increase in volume of the actual profile over the designed profile of each round. Apart from overbreak, underbreak may also occur in rock masses due to incorrect hole spacing, charging and other reasons. Factors influencing overbreak and underbreak may be grouped into two. The first group, geological factors, includes joint orientation, joint spacing, clay fillings and alteration, rock strength and ground stress effect. The second group, blasting factor, includes explosive type and powder factor, charge concentration, delay timing, perimeter blast hole pattern, drilling deviation, blast hole length and diameter and large hole cut (Kim and Moon, 2013). Parameters suggested to have influence on overbreak are compiled and presented in Table 9. The correctly designed blast is vital for controlling of overbreak and reducing support requirements. However, perimeter powder factor is highly responsible for overbreak and rock mass damage. Blast pattern of perimeter holes for different rock types is presented in Table 10. Charging of perimeter holes also contributes to overbreak in underground excavations. Overloading perimeter and buffer holes (the holes next to the perimeter), will damage rock beyond the opening perimeter and weaken the opening's stability, which increases scaling and cycle time (Mathis and Page, 1995). The amount of explosives in perimeter holes used in different blast hole lengths in various projects are presented in Table 11. The plot of amount of explosive used against the hole length shows a good correlation of 0.94 (see Fig. 12) . It can be used to control the charge required in a given blast hole length.
Table 8 Ground vibration monitoring at lake tap horizontal tunnel (Murthy and Dey, 2003). S No.
1 2 3 4 5 6 7 8 9 10
Distance (m)
43.1 44.5 45.9 47.3 48.7 50.5 52.3 53.7 55.1 57.1
Overbreak (%)
15.14 14.32 11.15 7.23 17.75 2.45
Q (kg)
V (mm/s)
Round-I
Round-II
Round-I
Round-II
18.75 12.75 6 – – 10.8 6.6 9 16.8 13.65
– – 7.7 7.2 5.25 – – 4.8 9 3.6
20.78 13.30 11.50 – – 12.22 8.86 12.76 11.73 11.13
– – 8.2 7.78 6.22 – – 6.65 5.86 1.83
Fig. 10. Blast pattern for lake tap tunnel, Koyna (Murthy and Dey, 2003). 314
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Average overbreak and the perimeter holes charging data were collected from four different tunnels driven in granitic rock masses of different rock classes and are shown in Table 12 (see Fig. 15).
Overbreak and explosive charges both depend on the rock classes. Results of regression analysis for overbreak (OB) and pull (P) prediction in varied rock classes are deduced as given in Eqs. (9) and (10).
OB = 1.6P
(9)
0.06Q + 15.6
(10)
P = 0.03Q + 0.022OB + 1.22
where OB = Overbreak, cm, Q = Rock mass quality value and P = Pull, m. 6. Results and discussion From the analysis of data compiled from various tunnelling cases the following key parameters emerge for tunnel blasting: 6.1. Pull versus rock mass quality The variation of pull with Q-values is shown in Fig. 13. It may be observed that pull is closely related to the rock mass quality. Higher Qvalues yield higher pulls in general. However, a significant scattering of values is observed which could be due to multi-parametric influence
Fig. 11. A graph of PPV vs Modified scale distance.
Table 9 Suggested models to control overbreak in tunnel blasting. Author(s)
Parameters
Techniques
Kim and Moon (2013)
Rock mass quality (classification), Deviation from designed contour, Contour hole spacing and burden, Look out angle, Contour holes charging
Pusch and Stanfors (1992)
Fracture characterization
Ibarra et al. (1996)
Q-value, Perimeter powder factor
Murthy and Dey (2003)
Maximum charge per delay, PPV
Drawing contour holes Drilling contour holes Charging Normal blasting Careful blasting Very careful blasting Rock classification Proper charging Controlled blasting in tunneling
Table 10 Blast pattern design for perimeter holes. Type of holes
RMR
< 20
20–40
41–60
61–80
81–100
Reference
Perimeter holes
Spacing (m) Burden (m) Rock type Spacing (m) Burden (m)
0.5–0.55 0.65 Soft 14Øh 1.2 * Spacing
0.55–0.6 0.70
0.6–0.65 0.75 Medium 15Øh 1.2 * Spacing
0.65–0.7 0.85
Max 0.75 Max 0.90 Hard 16Øh 1.2 * Spacing
Dey and Murthy (2011) Singh and Xavier (2005)
Note: Øh is blast hole diameter (m). Table 11 Charging scheme for perimeter holes in different tunnelling projects. Rock Class
Hole diameter (mm)
Hole length (m)
Charge diameter (mm)
Perimeter charge (kg)
Reference
Q > 10
42 45 42 45 45 45 45
4 3.5 3 3 3 2 1.5
32 32 + (6)17 32 32 + (5)17 32 + (5)17 25 25
1.10 0.975 0.80 0.875 0.875 0.75 0.60
Yang et al. (2014) Kim and Moon (2013) Chen et al. (2015) Kim and Moon (2013) Kim and Moon (2013) Murthy and Dey (2003) Murthy and Dey (2003)
0.1 < Q < 10
45 45 45 45 45
2.5 2 1.6 1.5 1.5
38 32 + 17 25 32 + 17 32 + 17
0.50 0.675 0.625 0.425 0.488
Chakraborty et al. (1998) Kim and Moon (2013) Murthy and Dey (2004) Kim and Moon (2013) Kim and Moon (2013)
Q < 0.1
43 43 43 45
2.3 1.7 1.7 1.2
34 34 34 32 + 17
0.772 0.772 0.872 0.325
Kuzu and Guclu (2009) Kuzu and Guclu (2009) Kuzu and Guclu (2009) Kim and Moon (2013)
315
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
(1) Rock breakage is more effective in hard and stable rocks due to better reflection of waves. (2) Percentage pull reduces due to rock collapse in weak rock (Q < 0.1) in comparison to hard rocks (Q > 0.1). (3) Method of tunnel support system as per stand up time. (4) The actual strength of rock mass is not adequately represented in all the cases. (5) Effective pull is dependent on area, width of opening, depth of drilling and cut. 6.2. Optimum charge versus Rockmass Quality Charging of perimeter holes with diameter of holes ranging from 42 to 45 mm and explosives having VOD ranging from 4000 to 6000 m/s is given in Table 13. Expected pull and charges used have been found to have controlling effect on the amount of overbreak obtained. A relation was developed to estimate the explosive charge requirements for controlling overbreak while attempting an optimum pull (Fig. 15). It may be observed that optimum pull (for minimum overbreak) lies in the range of 2.5–3 m and thus, allowable charge for minimum overbreak can be obtained as 0.8 kg for that pull. Considering the classification given in Table 2 and round length provided in various projects for different classes, the round length and sequence of excavation are suggested as shown in Table 14. The sequence of excavation for top heading and bench has been decided on the basis of the strength and stand up time of rock mass belonging to that class.
Fig. 12. Explosive amount required in blast hole. Table 12 Charge and overbreak in four tunnels (Kim and Moon, 2013). Q-value (used)
RMR class
Pull (m)
Charge (kg)
Overbreak (cm)
70 54 5.9 0.6 0.07
81–100 61–80 4–60 21–40 < 20
3.0 3.5 2.0 1.5 1.2
0.875 0.975 0.675 0.425 0.325
14.4 14.0 14.5 17.3 21.3
7. Conclusion Although, the round length has a major technical and economic impact in conventional tunnelling, no coherent procedure is available for its determination. In this study, determination of round length of excavation in varied rock classes was presented. The sequence of excavation, requirements for both top heading and bench were also addressed. Different classification systems were used to establish rock classes. As evident in this paper, characterization of the ground is the most important step towards the successful excavation and support for underground structures. On the other hand, characterization and determination of damage zone is to be carried out effectively for determination of support and also for safety of the working crew. Furthermore, overbreak control by means of optimum spacing and burden as well as charging of perimeter holes was suggested. Although,
Fig. 13. Variation of pull with Q-values.
such as charge quantities or distribution in the pattern variation. On the other hand, a plot of maximum pull defined in Table 2 shows a stepwise increase in goes high. This can be seen in Fig. 14. Possible iation in pull are attributed to:
blast holes and blast in each rock class as pull as the rock class reasons for such var-
Fig. 14. Increase in pull with rock classes. 316
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Fig. 15. A plot of overbreak vs pull and charging of perimeter holes. Table 13 Charging scheme for perimeter holes in varied rock classes. Rock class
Q < 0.1
Hole length (m) Charge quantity (kg)
1.2 0.33
0.1 < Q < 10 1.7 0.65
2.3 0.77
1.5 0.43–0.5
Q > 10 1.6 0.63
2 0.68
1.5 0.6
2 0.75
3 0.88
3.5 0.98
4 1.1
Table 14 Optimum round length and excavation sequences in different rock classes. Class
Round length (m)
Invert
Stand-up timea
NATM
Q
RMR
Top heading
Bench
I II
> 70 10–70
> 80 65–80
3–5 3–4
4.5 4.5
Invert is not required Invert is not required
10 years for 15 m span 6 months for 8 m span
III IV
4–10 1–4
58–65 47–58
2.5–3 2–2.5
4.0 3.5
Invert is not required Invert can be decided after measurements
2–4 days for 5 m span 5–10 h for 2.5 m span
V
0.11–1
29–47
1.5–2
2.5
Invert is required with round length of 13.5 m
2 h for 2.5 m span
VI VII VIII
0.03–0.11 0.015–0.03 0.008–0.015
20–29 15–20 10–15
1.25–1.5 1.0–1.25 <1
1.5 1.1 1.0
Invert is required with round length of 13.5 m as well as side drifting
– 30 min for 1 m span 30 min for 1 m span
a
Excavation sequence
Modified after Palmström, 1993 and Kockar and Akgun, 2003.
detailed study of the site conditions are required for a particular project, this study will provide useful information for the optimization of the excavation especially in design stage, so as to achieve the desired monthly drivage rates and help in timely completion of tunnelling projects.
Bolghonabai, R., Hossaini, M.F., Mohammadi, M., Nazem, A., 2015. On the selection of an appropriate excavation pattern for urban tunnels with big cross-section: a case study. Int. J. Min. Geo-Eng. 49 (2), 297–307. Chakraborty, A.K., Roy, P. Pal, Jethwa, J.L., Gupta, R.N., 1998. Blast performance in small tunnels - a critical evaluation in underground metal mines. Tunn. Undergr. Space Technol. 13 (3), 331–339. Chakraborty, A.K., Raina, A.K., Ramulu, M., Choudhury, P.B., Haldar, A., Sahoo, P., Bandopadhyay, C., 2004. Development of rational models for tunnel blast prediction based on a parametric study. Geotech. Geol. Eng. 22, 477–496. Chen, Chao Shi, Liu, Ya-ching, 2007. A methodology for evaluation and classification of rock mass quality on tunnel engineering. Tunn. Undergr. Space Technol. 22, 377–387. Chen, M., Lu, W.B., Yan, P., Hu, Y.G., 2015. Blasting excavation induced damage of surrounding rock masses in deep-buried tunnels. KSCE J. Civ. Eng. 1–10. Dey, Kaushik, Murthy, V.M.S.R., 2011. Delineating rockmass damage zones in blasting from in-field seismic velocity and peak particle velocity measurement. Int. J. Eng., Sci. Technol. 3 (2), 51–62. Holmberg, R., Persson, P.A., 1979. Design of tunnel perimeter blast hole patterns to prevent rock damage. In: Jones, J.M. (Ed.), Proceedings of the Tunneling’79. Institution of Mining and Metallurgy, London, UK.
Appendix A. Supplementary material Supplementary data to this article can be found online at https:// doi.org/10.1016/j.tust.2018.11.029. References Adhikari, G.R., Babu, A.R., Balachander, R., Gupta, R.N., 1999. On the application of rock mass quality for blasting in large underground chambers. Tunn. Undergr. Space Technol. 14 (3), 367–375.
317
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy Ibarra, J.A., Maerz, N.H., Franklin, J.A., 1996. Overbreak and underbreak in underground openings part 2: causes and implications. Geotech. Geol. Eng. 14, 325–340. Karahan, E., 2010. Design of Excavation and Support Systems for the Çubukbeli Tunnel in Antalya (Master of Science Thesis). Middle East Technical University, pp. 26–31. Kim, Y., Moon, H., 2013. Application of the guideline for overbreak control in granitic rock masses in Korean tunnels. Tunn. Undergr. Space Technol. 35, 67–77. Kockar, M.K., Akgun, H., 2003. Methodology for tunnel and portal support design in mixed limestone, schist and phyllite conditions: a case study in Turkey. Int. J. Rock Mech. Min. Sci. 40, 173–196. Kuzu, C., Guclu, E., 2009. The problem of human response to blast induced vibrations in tunnel construction and mitigation of vibration effects using cautious blasting in halfface blasting rounds. Tunn. Undergr. Space Technol. 24, 53–56. Li, Z., Huang, H., 1994. The calculation of stability of tunnels under the effects of seismic wave of explosions. In: Proceedings of the 26th Department of Defence Explosives Safety Seminar, USA, Department of Defence Explosives Safety Board. Mandal, S.K., Singh, S.K., 2009. Evaluating extent and causes of overbreak in tunnels. Tunn. Undergr. Space Technol. 24, 22–36. Mathis, J.I., Page, C.I., 1995. Drifting in very poor- experience and analysis. In: 101st Annual Northwest Mining Association Convention, Spokane, Washington, pp. 6–8. Murthy, V.M.S.R., Dey, Kaushik, 2003. Predicting overbreak from vibration monitoring in a lake tap tunnel - a success story. Fragblast 7 (3), 149–166. Murthy, V.M.S.R., Dey, Kaushik, 2004. Development of predictive models for controlling blast-induced overbreak in tunnels. J. Rock Mech. Tunnel. Technol. 10 (1), P31–P47. Palmström, A., 1993. The new Austrian tunnelling method. In: Conf. on Fjellsprengningsteknikk, Bergmekanikk, Geoteknikk, Oslo, pp. 31.1–31.20. Pusch, R., Stanfors, R., 1992. The zone of disturbance around blasted tunnel at depth. Int.
J. Rock Mech. Min. Sci. Geomech. Abstr. 29 (5), 447–456. Ryu, Chang-Ha, Sunwoo, Choon, Lee, Sang-Don, Choi, Hae-Moon, 2006. Suggestions of rock classification methods for blast design and application to tunnel blasting. Tunn. Undergr. Space Technol. 21 (3–4), 401–402. Singh, S.P., 2012. Influence of Geology Blast Damage. Laurentian University, Sudbury, Ontario CIM Bulletin. Singh, S.P., Xavier, Peter, 2005. Causes, impact and control of overbreak in underground excavations. Tunn. Undergr. Space Technol. 20, 63–71. Sjogren, B., Ofsthus, A., Sandberg, J., 1979. Seismic classification of rock mass qualities. Geophys. Prospect. 27, 409–442. Ulukan, B., Akcelik, N., Firat, Cetin. 1998 The Reduction in Efficiency and the Excavation Difficulty Depending On Rock Classes in Tunnels. General directorate of HighwayTurkey, http://www.kgm.gov.tr/SiteCollectionDocuments/KGMdocuments/ Baskanliklar/BaskanliklarTeknikArastirma/Yeni%20Klas%C3%B6r/Yay%C4% B1mlar/EXCAVATION_DIFFICULTY_ROCK_CLASSES_TUNNELS.pdf. Wu, C., Hao, H., 2006. Numerical prediction of rock mass damage due to accidental explosions in an underground ammunition storage chamber. Shock Waves 15 (1), 43–54. Yang, J.H., Lu, W.B., Zhao, Z.G., Yan, P., Chen, M., 2014. Safety distance for secondary shotcrete subjected to blasting vibration in Jinping-II deep-buried tunnels. Tunn. Undergr. Space Technol. 43, 123–132. Yilmaz, O., Unlu, T., 2014. An application of the modified Holmberg-Persson approach for tunnel blasting design. Tunn. Undergr. Space Technol. 43, 113–122. Yu, C.W., Chern, J.C., 2007. Expert system for drilling and blasting tunnel construction. Underground Space – the 4th Dimension of Metropolises. Taylor & Francis Group, London ISBN 978-0-415-40807-3.
318
Contents lists available at ScienceDirect
Tunnelling and Underground Space Technology journal homepage: www.elsevier.com/locate/tust
Optimising blast pulls and controlling blast-induced excavation damage zone in tunnelling through varied rock classes Amiri Hamis Saluma,
⁎,1
a b
T
, V.M.S.R. Murthyb
Former Postgraduate student at Indian Institute of Technology (Indian School of Mines), Dhanbad, India Professor & Head of Department, Department of Mining Engineering, Indian Institute of Technology (Indian School of Mines) Dhanbad, Jharkhand 826004, India
ARTICLE INFO
ABSTRACT
Keywords: Rock mass characterization Round length Excavation sequence Overbreak Explosive charge quantity
Tunnelling is the sole means of providing access for transportation, water conveyance in hydropower, mining of minerals, etc. Hydropower alone contributes to about 19% of the commercial energy needs of India and construction of hydroelectric power projects to meet these needs involves huge amounts of tunnelling through varied rock mass conditions. This necessitates identifying methods of tunnelling that are cost-effective, suited to varied geology and yet rapid to commission the projects in place of continuous excavation systems like TBM and Roadheader which are geology sensitive, costly and completely imported thus constraining their smooth adaptation. Mechanized drilling and advanced blasting techniques are the most often used methods of excavation of tunnels in India and in other countries for rapid and cost effective tunnelling, which depends, to a large extent, on the blast performance. Among the outcomes of any blasting operation, pull or advance achieved per blast and blast-induced excavation damage are of major concerns. It is essential to limit the blast-induced damage so as to control time and cost overruns in an underground project more so in varied geology. This paper discusses determination of optimum round lengths of excavation in varied rock classes as well as controlling overbreak in tunnelling. The sequence of excavation, requirements for both top heading and bench are also addressed. It has been observed from the past experiences that round lengths up to 5 m are practiced in rock class I and as the class improves to VII, the pull attempted reduces to about 1 m. Since charging of perimeter holes contributes to overbreak in underground excavations, a thorough analysis of the design of blasting pattern and scheme of charging for minimizing overbreak has been suggested. Characterization of ground through seismic imaging coupled with ground vibration monitoring has been suggested to control blast – induced rock damage and also arrive at optimum charges. As evident in this paper, characterization of the ground is the most important step towards rapid tunnelling.
1. Introduction Drilling and blasting method is generally inevitable for rock excavation activities in mining, quarrying and civil construction works. Use of explosive energy probably is the most widely used means of crushing rock, as well as the most cost-effective rock excavation method (Yilmaz and Unlu, 2014). Compatibility and feasibility to any sudden required alteration in dimension of excavation profile and/or geological constraints also adds to the popularity and suitability of this method over any other methods of excavation such as TBM’s, Road headers and Impact hammers (Mandal and Singh, 2009). However, performance of drilling and blasting method is mainly influenced by a number of factors that can be categorized as: Rock mass features, explosives characteristics and their distribution, blast design
and execution. For a given tunnel alignment, rock mass features cannot be changed but their knowledge facilitates the judicious selection of the drilling systems and patterns, explosive characteristics and the blast design parameters to obtain optimum pull and reduced excavation damage (Singh and Xavier, 2005). The tunnel blast performance is generally measured in terms of one or more than one of the following blast parameters: 1. Pull (face advance/depth of round), expressed in percent, 2. Specific charge (kg of explosive/m3 or t of yield), 3. Specific drilling (m of drilling/m3 or t of yield), or Detonator or hole factor (number of holes/m3 or t of yield), and 4. Blast-induced rock mass damage or overbreak or underbreak (Chakraborty et al., 2004).
Corresponding author. E-mail addresses: [email protected] (A.H. Salum), [email protected] (V.M.S.R. Murthy). 1 Currently at: Tunnel Engineer, Unitec Civil Consultant Ltd, P.O.Box 32507, Dar-es-Salaam, Tanzania. ⁎
https://doi.org/10.1016/j.tust.2018.11.029 Received 20 May 2016; Received in revised form 28 July 2017; Accepted 24 November 2018 Available online 02 January 2019 0886-7798/ © 2018 Elsevier Ltd. All rights reserved.
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
2. Rock mass characterization for tunnel blasting The main focus in tunnelling projects is to achieve longer pulls (up to 6 m) considering the large tunnel size available. Currently, pull achieved in different blasting practices ranges from 1.5 m to 4 m and is shown in Fig. 1. In order to achieve longer pull with minimum rock damage, there is a need to go for thorough rock mass characterization so as to develop a proper blast design in terms of blast geometry, pull length, powder factor and choosing the right type of explosive for controlling the blastinduced excavation damage zone (EDZ). The interaction of parameters is shown in Fig. 2. Rock mass is a heterogeneous material, a fact rarely considered and properly represented in blast design. In reality rock mass features have a controlling influence on the outcome of a blast (Singh and Xavier, 2005). The rock mass parameters which influence blast results fall into two groups. The first group is the intact rock properties and the second group is the discontinuity structure (Ryu et al., 2006). For the purpose of blast design, it is very important to consider the physico-mechanical including dynamic properties of rock due to the fact that under dynamic loading, the strength of rock is reported to vary with the rate of loading or strain rate. The dynamic properties of rocks to be considered during design are: Young’s modulus (E), Poisson’s ratio (ν), compressive strength (σc), dynamic tensile strength (σtd), P-wave velocity (Cp), and S-wave velocity (Cs). Other static properties include density of rock (ρr) and specific gravity (Y). Rock mass features like orientation of discontinuities, aperture of discontinuities, frequency of discontinuities and filling in the joints contribute significantly in producing damage to the perimeter of an excavation. Rock Quality Designation (RQD) also gives fair indication of the quality of rock mass (Singh and Xavier, 2005).
Fig. 1. Pull achieved in various drilling and blasting operations in tunnelling and mining.
2.1. Rock mass classification vis-à-vis tunnelling There are different classification systems used to describe the rock mass quality. Among these classification systems, Rock Mass Rating (RMR), Qsystem and New Austrian Tunnelling Methods (NATM) are widely used for excavation and support design in tunnelling. However, in this study the following classes and ranges as shown in Table 1 are adopted. In the table, ten classes of rocks based on NATM classification and their approximate range in Q-system and RMR are described. Behaviour of each class is also explained to give complete understanding of the rock class. Method of rock excavation, especially, by drilling and blasting varies depending on the quality of rock and the likelihood of rock damage. Typical drilling and blasting schemes for tunnelling through varied classes are shown in Fig. 3. Normal blasting can be used in class I and II with fairly good results. But rocks belonging to Class III through V require cautious blasting. Class VI and VII require pre support like forepoling, baby arch and low strength explosives. Excavation of rocks belonging to Class VIII and below, make use of very light blasting or mechanical excavation (Mathis and Page, 1995).
Fig. 2. Blasting system in tunnelling and interaction of parameters.
Two main objectives of any tunnel blasting operation are to get higher drivage rates and minimize damage to the surrounding rock mass. Attempts to get higher pull sometimes lead to roof rock damage, more so in case of weak and jointed rocks. In order to optimize blastinduced rock damage, rock mass characterization and assessment of the extent of rock damage are the pre requisites. Most of the existing criteria relate damage to ground vibrations resulting from dynamic stresses induced by blasting practices (Murthy and Dey, 2004). Thus, it is very important to characterize the rock mass for blasting purposes in order to achieve optimum results. This paper presents a methodology of characterizing the rock mass, various theoretical approaches to assist in achieving higher pull and minimize damage in tunnelling.
Table 1 Approximate relationship between NATM, Q-system and RMR system (Karahan, 2010). Class
NATM system term
Q-system
RMR
Remarks
I II III IV V VI VII VIII IX X
Stable Slightly Overbreaking Friable Very Friable Rolling Rock Bursting Squeezing Heavily Squeezing Flowing Swelling
> 70 10–70 4–10 1–4 0.11–1 0.03–0.11 0.015–0.03 0.008–0.015 0.002–0.008 < 0.002
> 80 65–80 58–65 47–58 29–47 20–29 15–20 10–15 5–10 <5
The rock mass is permanently stable without support A slight tendency of shallow overbreaks in the tunnel roof and in the upper portions of the sidewalls Overbreaks and loosening of the rock strata in tunnel roof and upper sidewalls if no support is installed in time Stand-up time and unsupported span are short Failure mechanisms such as spalling, buckling, shearing and rupture of the rock structure Rapid and significant movement of the rock mass into the cavity Limit the unsupported spans at arch and face Prior installation of forepoling or forepiling and shotcrete sealing of faces. The low cohesion requires a number of subdivisions
308
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
(a) Class I – II
(b) Class III – V
(d) Class VIII - X
(c ) Class VI – VII
Fig. 3. Illustration of different drilling and blasting techniques.
induced stresses, regional geology, structural geology and weak zone characteristics (Yu and Chern, 2007). It is possible to excavate full face for tunnels in Class I and II provided that the face area is within the reach of the drilling jumbo. But due to large cross section of some tunnels like motorway tunnels, staged excavation (Top heading and bench) is preferred in practice. However, in rock classes V and below multistage (Side drifting) is required for the stability of tunnels (Ulukan et al., 1998). Also an invert is usually provided for stability purpose (Chen and Liu, 2007). Yu and Chern (2007) have proposed a diagram that can be used as a basis for selection of excavation method (Fig. 4). Thus, the excavation method can be determined according to the given span and ratio of Uniaxial Compressive Strength (UCS) to vertical in situ stress. This concept was used in the design of excavation sequence in Niayesh Tunnel Project in Iran. This was the biggest tunnelling project in urban area in Middle East for its length, cross section area and step of the route. The tunnel project is mouth shaped, and is 13 m high and 18 m wide. Using Yu and Chern graph, it was revealed that Central diaphragm and side drifting methods are appropriate excavation patterns (Fig. 5). However, side drifting was preferred due to low settlement as observed from numerical modelling results of the two (Bolghonabai et al., 2015).
Fig. 4. Empirical determination of excavation sequences based on span size, unconfined compressive strength (UCS) and vertical stress on tunnel (Yu and Chern, 2007).
2.2. Determination of excavation method and sequences Generally, there is no simple rule to facilitate decision making about optimal selection of excavation method. However, some influencing factors on selection of excavation method can be mentioned considering rock mass properties including intact rock and joint (discontinuity) characteristics, shape and size of tunnel section, underground hydrology, in situ and
3. Determination of advance per round in tunnelling Traditionally, the advance per round is designed based on tunnel 309
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
(b) Central diaphragm (CD)
(a) Side drift
Fig. 5. Excavation stages selection for Niayesh Tunnel (Bolghonabai et al., 2015).
(a) Wedge cut in a small tunnel
(b) Parallel cut in small tunnel
Fig. 6. Variation in blasting parameters in different cuts practiced in tunnelling.
hydraulic radius in wedge cut and relief hole diameter in parallel cut. A wedge cut is more productive than parallel cut in small tunnels (Area < 10 m2) where the desired advance per round does not exceed 0.7 of the tunnel width. Here an advance rate of 0.4–0.7 times hydraulic radius can be obtained. (Chakraborty et al., 1998) On the other hand, the maximum advance per round in a parallel cut is influenced by the size of the relief holes but not the size of the opening. It is given by the following formula:
H = 0.15 + 34.1Ø
39.4Ø2
[Ø
0]
investigate the variation of pull with rock quality for eight classes (Table 2). Decision with regards to the round length must be verified with the RQD values as well as the sonic velocity of rocks. This is to ensure proper breaking of rock, by studying the impedance characteristics of the rocks so as to idealize the whole concept of energy transfer in the rock mass. Sjogren et al. (1979) studied RQD values based on sonic velocity data from various rock types and different geographical locations to analyze some of the relationships between seismic variables and their correlations with various physical properties of rock masses. Variation in cracks frequency and sonic velocity with RQD can be obtained directly from Fig. 7. RQD is an indicator of the quality of the rock mass. RQD of < 70% indicates that the rock mass will be more susceptible to blast damage and the RQD values < 50% would require close spacing, light loading and relief holes to produce acceptable results of overbreak (Singh and Xavier, 2005).
(1)
where H is the maximum advance per round, m, and Ø is the relief hole diameter, m. The parallel cut is less productive than the wedge cut in small size tunnels where the ratio of cut to opening area tends to one (Fig. 6). Higher charges and more number of blast holes are required in parallel cuts compared to a wedge cut. However, a parallel cut is more feasible in medium (Area = 10–35 m2) to large tunnels (Area > 35 m2) where the need of the higher quantity of charge and number of blast holes in the cut area can be neutralized by a higher ratio of the area between the tunnel opening and the cut section (Fig. 6). In parallel cut, usually a minimum of 25% of the volume of the initial cut should be removed by drilling relief holes (Adhikari et al., 1999). Pull designed in this way, does not take into account the rock quality which should be given due concern while designing tunnel blast. Data regarding pull were collected from about 40 case studies to
4. Blast vibrations and rock mass damage. Blasting damage in a rock mass is a result of the borehole strain developed by an explosive in a blast hole. According to Hooke’s law, the axial stress is related to the axial strain. This relationship can be expressed as:
= /E
(2)
where E is the elastic modulus of the material and σ and ε are the axial stress and strain respectively. 310
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Table 2 Blast round lengths in varied rock classes. S. No.
1 2 3 4 5
Excavation classes NATM CLASS Q system RMR Round length top heading (m) Round length of bench (m) Round length top heading (m) Round length top heading (m) Round length of Bench height (m) Round length top heading (m) Round length of Bench height (m) Round length for top heading (m) Round length for top heading (m)
Ref. I > 70 > 80 4.2 – <4
5 3.5
II 10–70 65–80 3 4.5 <3
4, 4.3, 3.2, 3 2.3, 2, 1.5
III 4.0–10 58–65 2.85 3.8 < 2.5 2–3 4 2–3 4 3 1.5, 2.5
IV 1.0–4.0 47–58 2.3 3.25 1.5–2 1.5–2.5 3.5 1.5–2 3–3.5 3 1.5
4.1. Rock damage assessment from peak particle velocities
(3)
Assuming brittle failure mode of the rock, a threshold of critical peak particle velocity PPVcrit which can be withstood by the rock before tensile failure can be computed from Eq. (4).
PPVcrit = (
t
VII 0.015–0.03 15–20 1.1 1.1 –
VIII 0.008–0.015 10–15
Ulukan et al. (1998) Chen and Liu (2007) Kockar and Akgun (2003)
0.75–1.25 <2
Kockar and Akgun (2003) From various projects
PPV (mm/s)
Tensile stress (MPa)
Strain energy (J/ Kg)
Typical effect in hard Scandinavian bedrock
700 1000 2500 5000 12,000
8.7 12.5 31.2 62.4 187
0.25 0.5 3.1 12.5 112.5
Incipient swelling Incipient damage Fragmentation Good fragmentation Crushing
(4)
Cp)/ E
Several researchers investigated the ground vibration levels during blasting in tunnels, quarrying and mines and suggested different threshold levels (PPV) for rock mass damage. The rock damage criteria suggested are presented in Table 4. Li and Huang (Wu and Hao, 2006) discussed PPV damage criterion for rock tunnels with the following definition: slight damage (initial cracking); medium damage (partial collapse); serious damage (largearea tunnel collapse). The respective PPV for the various types of rock mass are shown in Table 5. It may be observed that the average tensile strength values appear to relate well with the allowable threshold levels and this may be due to the breakage of rock being effected in reflection mode taking tensile strength as the key weakness. The relation is shown in Fig. 8. Thus, for a given tensile strength of rock the damage threshold levels (PPV) can be fixed and accordingly the charges can be controlled. On the other hand Singh (2012) suggested critical PPV values for different rock masses based on Compressive strength properties as given in Table 6. Either of tensile or compressive strength can be used to predict
where σt is the tensile strength. In addition, the vibration level above which some damage may be expected PPVmax can be estimated from Eq. (5).
PPVmax = 1.2[ t /(Cp
VI 0.03–0.11 20–29 1.15 1.5 0.8–1.2
Table 3 Effect of PPV levels in Scandinavian bedrock (Wu and Hao, 2006).
The peak strain in the rock mass can also be related to the PPV and compressional elastic wave velocity Cp as given by:
= PPV / Cp
V 0.11–1.0 29–47 – – 1–1.5 1.5–2 2.5 1.25–1.5 <3 2.5 1.7, 2, 2.3
r )]
(5)
where ρr is the density of rock in kg/m3 (Yilmaz and Unlu, 2014). Eqs. (3) and (4) can be used for determination of critical strain in rock mass of certain P-wave velocity, which is a characteristic of that medium. Knowledge of this critical strain will be used as a criterion for deciding the optimum or maximum pull for rock mass having similar properties. In addition to that, Eq. (4) is very important in determination of damage depth hence deciding the support requirement. Holmberg and Persson (1979) determined the limiting PPV based on new crack formation in the surrounding rock mass and they found that the critical peak particle velocity for new crack formation is between 700 and 1000 mm/s for Swedish hard igneous rocks (Holmberg and Persson, 1979; Wu and Hao, 2006). Table 3 shows the observed typical effects that would occur when the peak particle velocities vary in Scandinavian bedrock.
Fig. 7. Variation of cracks frequency and P-wave velocity with RQD. 311
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Table 4 Suggested damage threshold for rock mass damage (Murthy and Dey, 2003). Model (Year)
Damage criteria
Edwards and Northwood (1960), Langefors et al. (1973), Nicholls et al. (1971) Bauer and Calder (1970)
Langefors and Kihlstrom (1973) Holmberg and Persson (1979) Oriard (1982) Rustan (1985) Meyer and Dunn (1995) Bogdanhoff (1996) Murthy and Dey (2002) Dey (2004)
PPV (mm/s)
Damage description
< 50 < 254 254–635 635–2540 > 2540 305 610 700–1000 > 635 300–900 1000–3000 300 600 2000–2500 2050 700–1300
Low probability of structural damage to residential buildings No fracturing of intact rock Minor tensile slabbing Strong tensile slabbing and radial cracking Break up of rock mass Results in fall of rocks Results in formation of new cracks Rock mass damage Rock mass damage Smooth blasting Rock damage threshold Minor damage Damage threshold Damage threshold Threshold for overbreak in compact basalt Threshold for overbreak
Tunnelling Mining
Tunnelling Tunnelling Tunnelling Tunnelling Mining Tunnelling Tunnelling Metal mines
4.2. Blast-induced damage assessment using seismic imaging and PPV measurement
Table 5 Rock tunnel damage criteria (modified after Li and Huang (1994)). Average
Application
Peak particle velocity (mm/s)
Compressive strength (MPa)
Tensile strength (MPa)
No damage
Slight damage
Medium damage
Serious damage
92.5 145 190 70 130
2.75 4.25 5.4 2.1 3.95
270 310 360 290 350
540 620 720 580 700
820 960 1110 900 1070
1530 1780 2090 1670 1990
Blast-induced rock damage has been related with peak particle velocity (PPV) by many researchers and a mathematical formula has been proposed for assessing the extent of damage zone by extending the formula proposed by Holmberg and Persson (1979) for the near-field vibration approximation (Murthy and Dey, 2003). This model gives damage envelope when plotted in space (x, y). To test the acceptability of the proposed model, a crater blast experiment has been carried out in the bench of a surface mine with 1 m long and 32 mm dia drill holes, at Nuasahi chromite surface mine in Orissa. The holes have been blasted with 250 g (0.4 m) explosive. Vibration monitoring has been carried out close to blast site to establish the vibration predictor. Seismic imaging has been carried out at the blast site before and after the blast. Analysis of the seismic images (pre and post blast) has been done to decipher the
critical PPV with satisfactory results, but authors of this paper would prefer using tensile strength as the same can be used in Tezuka’s equation to calculate PPV and strain.
Fig. 8. Effect of rock properties on threshold levels of damage (modified after Li and Huang, 2008). 312
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Trial blasts were conducted and vibration monitoring was done for establishing the ground vibration attenuation equation. For both the rounds of blast conducted, the PPV was measured for varying maximum charge per delay (W) and radial distance (R). Table 8 presents the details of data generated. For minimizing the blast-induced vibration and related rock damage, the blasts were executed in two rounds. In Round-I, holes up to spring level were blasted and the remaining holes were blasted in Round-II, after necessary dressing (Fig. 10).
Table 6 Critical velocities for different types of rocks. Rock type
Hard Medium soft
Properties Specific gravity
Uniaxial compressive strength (MPa)
> 2.7 > 2.5 > 2.3
> 240 100–150 < 50
Critical PPV (mm/ s)
1200–2000 700–1000 < 400
5.2. Determination of PPV predictor equations
extent of rock damage. Using the proposed model, the damage envelope has been computed. Comparing the damage envelopes obtained from the proposed model and seismic imaging, it has been found that they are in close agreement. Thus, it can be inferred that the suggested PPV based model holds good to decipher the damage zone. Apart from this, it has been found that the damage zones predicted from seismic images were 2–30 times larger than the physically measured crater volumes. This probably includes the cracked as well as incipient cracks zone. Seismic imaging technique has, thus, been found suitable for determining the damage extent with considerable accuracy. The pre and post blast Pwave velocity values were obtained and are shown in Fig. 9(a) and (b) respectively. It may be observed that this method can help assess the damage zone surrounding the blast hole there by optimizing the charges that can be exploded to control the overbreak.
Using data presented in Table 10, ground vibration predictor equations were established between PPV and scaled distance (SD). For Round-I, the scaled distance was given by the formula,
SD =
3
R W
(6)
Therefore, the PPV predictor equation was given as:
V = 514.98 ×
3
R W
1.2
,
Correlation coefficient r = 0.87
(7)
where V = PPV (mm/s) R = Distance (m) W = Maximum charge/delay (kg) In case of Round-II, the scaled distance was modified by introducing the stiffness factor, defined as the ratio of burden to hole depth, whose value is 0.6. Hence the best fit ground vibration equation obtained is:
5. Field investigations, data collection and analysis Blast design and ground vibration monitoring were carried out in Koyna Hydel Project by the authors and the same has been described below.
V = 93.95 ×
5.1. Blast design and vibration data monitoring
R 1 W 3 + SF
1.24
,
correlation coefficient R = 0.80
(8)
Critical velocity is given by Vcr = = 0.635 m/s. The value obtained is slightly less than that proposed by Singh and Holmberg and Persson (0.7 m/s). The damage distance obtained is 0.25 m as shown in Fig. 11. t × VP E
A pull of 1.5 m was chosen for considering cycle optimization and other factors such as lack of advanced initiation systems. The intact properties of rock are presented in Table 7.
a) Pre-blast seismic image
b) Post blast seismic image
Fig. 9. P-wave velocity profile surrounding a blast hole for blast-induced damage assessment (Dey and Murthy, 2011). 313
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
5.3. Overbreak control in tunnelling
Table 7 Important rock properties of lake tap horizontal tunnel, Koyna Project, India (Murthy and Dey, 2003). Rock characteristics
Value
Value taken in this study
Q-value P-wave velocity (m/s) UCS (MPa) Density (g/cm3) Tensile strength (MPa) Young modulus (MPa)
15–21 1970–5200 13–95 2.16–2.97 1.2–8.1 33,000–66,330
18 5200 95 2.97 8.1 66,330
Overbreak may be defined as the percentage increase in volume of the actual profile over the designed profile of each round. Apart from overbreak, underbreak may also occur in rock masses due to incorrect hole spacing, charging and other reasons. Factors influencing overbreak and underbreak may be grouped into two. The first group, geological factors, includes joint orientation, joint spacing, clay fillings and alteration, rock strength and ground stress effect. The second group, blasting factor, includes explosive type and powder factor, charge concentration, delay timing, perimeter blast hole pattern, drilling deviation, blast hole length and diameter and large hole cut (Kim and Moon, 2013). Parameters suggested to have influence on overbreak are compiled and presented in Table 9. The correctly designed blast is vital for controlling of overbreak and reducing support requirements. However, perimeter powder factor is highly responsible for overbreak and rock mass damage. Blast pattern of perimeter holes for different rock types is presented in Table 10. Charging of perimeter holes also contributes to overbreak in underground excavations. Overloading perimeter and buffer holes (the holes next to the perimeter), will damage rock beyond the opening perimeter and weaken the opening's stability, which increases scaling and cycle time (Mathis and Page, 1995). The amount of explosives in perimeter holes used in different blast hole lengths in various projects are presented in Table 11. The plot of amount of explosive used against the hole length shows a good correlation of 0.94 (see Fig. 12) . It can be used to control the charge required in a given blast hole length.
Table 8 Ground vibration monitoring at lake tap horizontal tunnel (Murthy and Dey, 2003). S No.
1 2 3 4 5 6 7 8 9 10
Distance (m)
43.1 44.5 45.9 47.3 48.7 50.5 52.3 53.7 55.1 57.1
Overbreak (%)
15.14 14.32 11.15 7.23 17.75 2.45
Q (kg)
V (mm/s)
Round-I
Round-II
Round-I
Round-II
18.75 12.75 6 – – 10.8 6.6 9 16.8 13.65
– – 7.7 7.2 5.25 – – 4.8 9 3.6
20.78 13.30 11.50 – – 12.22 8.86 12.76 11.73 11.13
– – 8.2 7.78 6.22 – – 6.65 5.86 1.83
Fig. 10. Blast pattern for lake tap tunnel, Koyna (Murthy and Dey, 2003). 314
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Average overbreak and the perimeter holes charging data were collected from four different tunnels driven in granitic rock masses of different rock classes and are shown in Table 12 (see Fig. 15).
Overbreak and explosive charges both depend on the rock classes. Results of regression analysis for overbreak (OB) and pull (P) prediction in varied rock classes are deduced as given in Eqs. (9) and (10).
OB = 1.6P
(9)
0.06Q + 15.6
(10)
P = 0.03Q + 0.022OB + 1.22
where OB = Overbreak, cm, Q = Rock mass quality value and P = Pull, m. 6. Results and discussion From the analysis of data compiled from various tunnelling cases the following key parameters emerge for tunnel blasting: 6.1. Pull versus rock mass quality The variation of pull with Q-values is shown in Fig. 13. It may be observed that pull is closely related to the rock mass quality. Higher Qvalues yield higher pulls in general. However, a significant scattering of values is observed which could be due to multi-parametric influence
Fig. 11. A graph of PPV vs Modified scale distance.
Table 9 Suggested models to control overbreak in tunnel blasting. Author(s)
Parameters
Techniques
Kim and Moon (2013)
Rock mass quality (classification), Deviation from designed contour, Contour hole spacing and burden, Look out angle, Contour holes charging
Pusch and Stanfors (1992)
Fracture characterization
Ibarra et al. (1996)
Q-value, Perimeter powder factor
Murthy and Dey (2003)
Maximum charge per delay, PPV
Drawing contour holes Drilling contour holes Charging Normal blasting Careful blasting Very careful blasting Rock classification Proper charging Controlled blasting in tunneling
Table 10 Blast pattern design for perimeter holes. Type of holes
RMR
< 20
20–40
41–60
61–80
81–100
Reference
Perimeter holes
Spacing (m) Burden (m) Rock type Spacing (m) Burden (m)
0.5–0.55 0.65 Soft 14Øh 1.2 * Spacing
0.55–0.6 0.70
0.6–0.65 0.75 Medium 15Øh 1.2 * Spacing
0.65–0.7 0.85
Max 0.75 Max 0.90 Hard 16Øh 1.2 * Spacing
Dey and Murthy (2011) Singh and Xavier (2005)
Note: Øh is blast hole diameter (m). Table 11 Charging scheme for perimeter holes in different tunnelling projects. Rock Class
Hole diameter (mm)
Hole length (m)
Charge diameter (mm)
Perimeter charge (kg)
Reference
Q > 10
42 45 42 45 45 45 45
4 3.5 3 3 3 2 1.5
32 32 + (6)17 32 32 + (5)17 32 + (5)17 25 25
1.10 0.975 0.80 0.875 0.875 0.75 0.60
Yang et al. (2014) Kim and Moon (2013) Chen et al. (2015) Kim and Moon (2013) Kim and Moon (2013) Murthy and Dey (2003) Murthy and Dey (2003)
0.1 < Q < 10
45 45 45 45 45
2.5 2 1.6 1.5 1.5
38 32 + 17 25 32 + 17 32 + 17
0.50 0.675 0.625 0.425 0.488
Chakraborty et al. (1998) Kim and Moon (2013) Murthy and Dey (2004) Kim and Moon (2013) Kim and Moon (2013)
Q < 0.1
43 43 43 45
2.3 1.7 1.7 1.2
34 34 34 32 + 17
0.772 0.772 0.872 0.325
Kuzu and Guclu (2009) Kuzu and Guclu (2009) Kuzu and Guclu (2009) Kim and Moon (2013)
315
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
(1) Rock breakage is more effective in hard and stable rocks due to better reflection of waves. (2) Percentage pull reduces due to rock collapse in weak rock (Q < 0.1) in comparison to hard rocks (Q > 0.1). (3) Method of tunnel support system as per stand up time. (4) The actual strength of rock mass is not adequately represented in all the cases. (5) Effective pull is dependent on area, width of opening, depth of drilling and cut. 6.2. Optimum charge versus Rockmass Quality Charging of perimeter holes with diameter of holes ranging from 42 to 45 mm and explosives having VOD ranging from 4000 to 6000 m/s is given in Table 13. Expected pull and charges used have been found to have controlling effect on the amount of overbreak obtained. A relation was developed to estimate the explosive charge requirements for controlling overbreak while attempting an optimum pull (Fig. 15). It may be observed that optimum pull (for minimum overbreak) lies in the range of 2.5–3 m and thus, allowable charge for minimum overbreak can be obtained as 0.8 kg for that pull. Considering the classification given in Table 2 and round length provided in various projects for different classes, the round length and sequence of excavation are suggested as shown in Table 14. The sequence of excavation for top heading and bench has been decided on the basis of the strength and stand up time of rock mass belonging to that class.
Fig. 12. Explosive amount required in blast hole. Table 12 Charge and overbreak in four tunnels (Kim and Moon, 2013). Q-value (used)
RMR class
Pull (m)
Charge (kg)
Overbreak (cm)
70 54 5.9 0.6 0.07
81–100 61–80 4–60 21–40 < 20
3.0 3.5 2.0 1.5 1.2
0.875 0.975 0.675 0.425 0.325
14.4 14.0 14.5 17.3 21.3
7. Conclusion Although, the round length has a major technical and economic impact in conventional tunnelling, no coherent procedure is available for its determination. In this study, determination of round length of excavation in varied rock classes was presented. The sequence of excavation, requirements for both top heading and bench were also addressed. Different classification systems were used to establish rock classes. As evident in this paper, characterization of the ground is the most important step towards the successful excavation and support for underground structures. On the other hand, characterization and determination of damage zone is to be carried out effectively for determination of support and also for safety of the working crew. Furthermore, overbreak control by means of optimum spacing and burden as well as charging of perimeter holes was suggested. Although,
Fig. 13. Variation of pull with Q-values.
such as charge quantities or distribution in the pattern variation. On the other hand, a plot of maximum pull defined in Table 2 shows a stepwise increase in goes high. This can be seen in Fig. 14. Possible iation in pull are attributed to:
blast holes and blast in each rock class as pull as the rock class reasons for such var-
Fig. 14. Increase in pull with rock classes. 316
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy
Fig. 15. A plot of overbreak vs pull and charging of perimeter holes. Table 13 Charging scheme for perimeter holes in varied rock classes. Rock class
Q < 0.1
Hole length (m) Charge quantity (kg)
1.2 0.33
0.1 < Q < 10 1.7 0.65
2.3 0.77
1.5 0.43–0.5
Q > 10 1.6 0.63
2 0.68
1.5 0.6
2 0.75
3 0.88
3.5 0.98
4 1.1
Table 14 Optimum round length and excavation sequences in different rock classes. Class
Round length (m)
Invert
Stand-up timea
NATM
Q
RMR
Top heading
Bench
I II
> 70 10–70
> 80 65–80
3–5 3–4
4.5 4.5
Invert is not required Invert is not required
10 years for 15 m span 6 months for 8 m span
III IV
4–10 1–4
58–65 47–58
2.5–3 2–2.5
4.0 3.5
Invert is not required Invert can be decided after measurements
2–4 days for 5 m span 5–10 h for 2.5 m span
V
0.11–1
29–47
1.5–2
2.5
Invert is required with round length of 13.5 m
2 h for 2.5 m span
VI VII VIII
0.03–0.11 0.015–0.03 0.008–0.015
20–29 15–20 10–15
1.25–1.5 1.0–1.25 <1
1.5 1.1 1.0
Invert is required with round length of 13.5 m as well as side drifting
– 30 min for 1 m span 30 min for 1 m span
a
Excavation sequence
Modified after Palmström, 1993 and Kockar and Akgun, 2003.
detailed study of the site conditions are required for a particular project, this study will provide useful information for the optimization of the excavation especially in design stage, so as to achieve the desired monthly drivage rates and help in timely completion of tunnelling projects.
Bolghonabai, R., Hossaini, M.F., Mohammadi, M., Nazem, A., 2015. On the selection of an appropriate excavation pattern for urban tunnels with big cross-section: a case study. Int. J. Min. Geo-Eng. 49 (2), 297–307. Chakraborty, A.K., Roy, P. Pal, Jethwa, J.L., Gupta, R.N., 1998. Blast performance in small tunnels - a critical evaluation in underground metal mines. Tunn. Undergr. Space Technol. 13 (3), 331–339. Chakraborty, A.K., Raina, A.K., Ramulu, M., Choudhury, P.B., Haldar, A., Sahoo, P., Bandopadhyay, C., 2004. Development of rational models for tunnel blast prediction based on a parametric study. Geotech. Geol. Eng. 22, 477–496. Chen, Chao Shi, Liu, Ya-ching, 2007. A methodology for evaluation and classification of rock mass quality on tunnel engineering. Tunn. Undergr. Space Technol. 22, 377–387. Chen, M., Lu, W.B., Yan, P., Hu, Y.G., 2015. Blasting excavation induced damage of surrounding rock masses in deep-buried tunnels. KSCE J. Civ. Eng. 1–10. Dey, Kaushik, Murthy, V.M.S.R., 2011. Delineating rockmass damage zones in blasting from in-field seismic velocity and peak particle velocity measurement. Int. J. Eng., Sci. Technol. 3 (2), 51–62. Holmberg, R., Persson, P.A., 1979. Design of tunnel perimeter blast hole patterns to prevent rock damage. In: Jones, J.M. (Ed.), Proceedings of the Tunneling’79. Institution of Mining and Metallurgy, London, UK.
Appendix A. Supplementary material Supplementary data to this article can be found online at https:// doi.org/10.1016/j.tust.2018.11.029. References Adhikari, G.R., Babu, A.R., Balachander, R., Gupta, R.N., 1999. On the application of rock mass quality for blasting in large underground chambers. Tunn. Undergr. Space Technol. 14 (3), 367–375.
317
Tunnelling and Underground Space Technology 85 (2019) 307–318
A.H. Salum, V.M.S.R. Murthy Ibarra, J.A., Maerz, N.H., Franklin, J.A., 1996. Overbreak and underbreak in underground openings part 2: causes and implications. Geotech. Geol. Eng. 14, 325–340. Karahan, E., 2010. Design of Excavation and Support Systems for the Çubukbeli Tunnel in Antalya (Master of Science Thesis). Middle East Technical University, pp. 26–31. Kim, Y., Moon, H., 2013. Application of the guideline for overbreak control in granitic rock masses in Korean tunnels. Tunn. Undergr. Space Technol. 35, 67–77. Kockar, M.K., Akgun, H., 2003. Methodology for tunnel and portal support design in mixed limestone, schist and phyllite conditions: a case study in Turkey. Int. J. Rock Mech. Min. Sci. 40, 173–196. Kuzu, C., Guclu, E., 2009. The problem of human response to blast induced vibrations in tunnel construction and mitigation of vibration effects using cautious blasting in halfface blasting rounds. Tunn. Undergr. Space Technol. 24, 53–56. Li, Z., Huang, H., 1994. The calculation of stability of tunnels under the effects of seismic wave of explosions. In: Proceedings of the 26th Department of Defence Explosives Safety Seminar, USA, Department of Defence Explosives Safety Board. Mandal, S.K., Singh, S.K., 2009. Evaluating extent and causes of overbreak in tunnels. Tunn. Undergr. Space Technol. 24, 22–36. Mathis, J.I., Page, C.I., 1995. Drifting in very poor- experience and analysis. In: 101st Annual Northwest Mining Association Convention, Spokane, Washington, pp. 6–8. Murthy, V.M.S.R., Dey, Kaushik, 2003. Predicting overbreak from vibration monitoring in a lake tap tunnel - a success story. Fragblast 7 (3), 149–166. Murthy, V.M.S.R., Dey, Kaushik, 2004. Development of predictive models for controlling blast-induced overbreak in tunnels. J. Rock Mech. Tunnel. Technol. 10 (1), P31–P47. Palmström, A., 1993. The new Austrian tunnelling method. In: Conf. on Fjellsprengningsteknikk, Bergmekanikk, Geoteknikk, Oslo, pp. 31.1–31.20. Pusch, R., Stanfors, R., 1992. The zone of disturbance around blasted tunnel at depth. Int.
J. Rock Mech. Min. Sci. Geomech. Abstr. 29 (5), 447–456. Ryu, Chang-Ha, Sunwoo, Choon, Lee, Sang-Don, Choi, Hae-Moon, 2006. Suggestions of rock classification methods for blast design and application to tunnel blasting. Tunn. Undergr. Space Technol. 21 (3–4), 401–402. Singh, S.P., 2012. Influence of Geology Blast Damage. Laurentian University, Sudbury, Ontario CIM Bulletin. Singh, S.P., Xavier, Peter, 2005. Causes, impact and control of overbreak in underground excavations. Tunn. Undergr. Space Technol. 20, 63–71. Sjogren, B., Ofsthus, A., Sandberg, J., 1979. Seismic classification of rock mass qualities. Geophys. Prospect. 27, 409–442. Ulukan, B., Akcelik, N., Firat, Cetin. 1998 The Reduction in Efficiency and the Excavation Difficulty Depending On Rock Classes in Tunnels. General directorate of HighwayTurkey, http://www.kgm.gov.tr/SiteCollectionDocuments/KGMdocuments/ Baskanliklar/BaskanliklarTeknikArastirma/Yeni%20Klas%C3%B6r/Yay%C4% B1mlar/EXCAVATION_DIFFICULTY_ROCK_CLASSES_TUNNELS.pdf. Wu, C., Hao, H., 2006. Numerical prediction of rock mass damage due to accidental explosions in an underground ammunition storage chamber. Shock Waves 15 (1), 43–54. Yang, J.H., Lu, W.B., Zhao, Z.G., Yan, P., Chen, M., 2014. Safety distance for secondary shotcrete subjected to blasting vibration in Jinping-II deep-buried tunnels. Tunn. Undergr. Space Technol. 43, 123–132. Yilmaz, O., Unlu, T., 2014. An application of the modified Holmberg-Persson approach for tunnel blasting design. Tunn. Undergr. Space Technol. 43, 113–122. Yu, C.W., Chern, J.C., 2007. Expert system for drilling and blasting tunnel construction. Underground Space – the 4th Dimension of Metropolises. Taylor & Francis Group, London ISBN 978-0-415-40807-3.
318