Blast induced rock mass damage around tunnels

Blast induced rock mass damage around tunnels

Tunnelling and Underground Space Technology 71 (2018) 149–158 Contents lists available at ScienceDirect Tunnelling and Underground Space Technology ...

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Tunnelling and Underground Space Technology 71 (2018) 149–158

Contents lists available at ScienceDirect

Tunnelling and Underground Space Technology journal homepage: www.elsevier.com/locate/tust

Blast induced rock mass damage around tunnels a,⁎

b

b

a

MARK a

H.K. Verma , N.K. Samadhiya , M. Singh , R.K. Goel , P.K. Singh a b

CSIR-Central Institute of Mining and Fuel Research, Barwa Road, Dhanbad 826015, India Indian Institute of Technology Roorkee, Roorkee 247667, India

A R T I C L E I N F O

A B S T R A C T

Keywords: Rock mass damage Blasting Maximum charge per delay Vibration Attenuation characteristics Perimeter charge factor

Drilling and blasting is a preferred method of rock excavation world-wide due to low initial investment, cheap explosive energy, easy acceptability among the blasting engineers and, possibility to deal with different shapes and sizes of openings. Although, drill and blast method has witnessed significant technological advancements, it has inherent disadvantage of deteriorating surrounding rock mass due to development of network of fine cracks in it leading to safety and stability problems. The damage in the peripheral rock mass culminates in the form of overbreak and damaged zone beyond overbreak. In some cases the projects cost has increased more than 15% because of overbreak. Although significant efforts have been made to assess damage to the surrounding rock mass using different methods, the solution based on easily available site parameters is still lacking. Authors have carried out field investigations at five different tunnels located in Himalaya, India to study blast induced damage for wide range of rock mass quality Q values (0.03–17.8). In addition to Q, specific charge, perimeter charge factor, maximum charge per delay, advancement and confinement factors have also been used. Data sets of 113 experimental blasts are collected from the five tunnel sites. All the parameters, easily available to the site engineers, have been used for developing an empirical correlation to estimate the rock mass damage around the tunnel, which is discussed in the paper. The proposed empirical correlation has been validated using ultrasonic tests on rock core samples obtained from one of the experimental location.

1. Introduction Rock excavation using drill and blast method (DBM) is commonly used in mining, quarrying and tunnelling world-wide. The drill and blast method is economical as compared to other mechanical methods utilizing rock breakers, tunnel boring machines and road headers especially with regards to tunnels excavation in varying geological conditions. Low initial investment, cheap explosive energy, easy acceptability among the blasting engineers, possibility to deal with different shapes and sizes of openings and reasonably faster advance rate in a suitable geotechnical mining condition collectively make DBM preferred method of rock excavation (Innaurato et al., 1998; Murthy and Dey, 2003 and Verma et al., 2015). The drill and blast method has witnessed considerable technological advancements particularly in the area of explosives, initiating devices, automation in drilling techniques and blast designs (Dey and Murthy, 2011). Despite the technological advancement, DBM has the inherent disadvantage of damaging the surrounding rock mass resulting in the development of network of blast-induced cracks in the surrounding rock masses leading to safety and stability problems. Blasting for underground excavation and tunnelling are difficult



Corresponding author. E-mail address: [email protected] (H.K. Verma).

http://dx.doi.org/10.1016/j.tust.2017.08.019 Received 9 March 2017; Received in revised form 27 July 2017; Accepted 16 August 2017 0886-7798/ © 2017 Published by Elsevier Ltd.

operations compared to open pit excavation due to lack of free face (Gupta et al., 1988; Adhikari and Babu, 1994 and Murthy and Dey, 2002). Practicing engineers attempt to achieve faster advancement in tunnel and underground excavation by employing drill jumbos. Such drill machine significantly reduces drilling time with improved accuracy. Faster advancement rate using higher amount of explosives leads to greater extent of blast induced rock mass damage (Murthy and Dey, 2003). Perimeter blasting techniques, such as smooth blasting (Holmberg and Persson, 1980) are commonly used to minimize damage to surrounding rock mass beyond the designed profile of tunnel. Despite the improvement in blasting techniques, rock mass damage is still inevitable and is evident in the form of increased support cost, slow tunnel advancement, unstable rock mass, prolonged incubation period of the projects and enhanced post-construction tunnel maintenance cost. Various researchers have studied and given emphasis on determining the extent of unwanted damage induced by blasting beyond the desired perimeter of the tunnel. The significance and importance of this damage have been deliberated by various researchers (Langefors and Kihlstrom, 1963; Bauer and Calder, 1978; Oriad, 1982; MacKown, 1986; Singh, 1993; Scoble et al., 1997; Backblom and Martin, 1999;

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Fig. 1. Blast induced Rock Mass Damage Zone Tunnels (Adapted from Singh and Xavier, 2005).

smallest disturbance to the rock mass may have significant implications due to possible percolation of contaminants along the fine cracks. Rock mass damage in mining and dimensional stone industries causes ore dilution. In tunnels too, rock mass damage has significant influence on cost and safety aspects. Extent and characterization of damaged zone pertaining to design and development of high level nuclear waste disposal repositories have been extensively studied (Martino and Chandler, 2004 and Hudson et al., 2009; Waltona et al., 2015). Daemen (2011) have emphasized on the importance of excavation damage zone (EDZ) assessment in design of nuclear waste repositories, especially at locations where permanent seals are to be installed. Importance to the blast induced rock mass damage in underground mining and tunnelling has, however, received relatively less attention (Scoble et al., 1997). In rock mass damage studies pertaining to tunnels, overbreak zone alone has been considered invariably as damage zone, whereas it has been found that the damage by blast extends beyond overbreak zone and plays vital role in the stability of underground structures in the long-term sometimes. Mandal and Singh (2009) suggested that the damaged zone beyond overbreak zone should be considered in the design of the tunnel support systems. Although significant efforts have been made to assess damage to the surrounding rock mass using different methods, the solution based on easily available site parameters is still missing. Review of available literature reveals that the results obtained from various blast induced damage estimation methods are inconsistent (Raina et al., 2000). Most of the methods are based on few cases and applicable to limited range of rock types (Raina et al., 2000). The evaluation of rock mass damage from the surface geometry of the tunnel can be done by various methods such as manual measurements, standard surveying, laser surveying with reflectors, photographic sectioning and light sectioning methods. The limitations of these methods are that they are too subjective, manually intensive, time-consuming and often provide information only for a limited sections (Warneke et al., 2007). Moreover, in some cases this will provide the information about the overbreak and not the extent of damage in peripheral rock mass. Some of the damage prediction models are based on laboratory investigations only wherein a single hole blast is considered. In actual field conditions because of number of holes, the quantity of explosive and interaction of different parameters make the problem complex and hence the simplistic laboratory scale study may not be able to

Raina et al., 2000; Ouchterlony et al., 2002; Singh and Xavier, 2005; Warneke et al., 2007; Ramulu et al., 2009 and Fu et al., 2014). Damage around an opening in underground has been described by using terminology such as blast induced rock mass damage (BIRD), blast induced damage (BID), excavation damage zone (EDZ), rock mass damage zone (RMD) etc. Blast induced rock mass damage zone surrounding an underground opening consists of overbreak zone (failed zone), damaged zone and a disturbed zone. In the present research work, the definition and significance of the three zones are as discussed below and shown in Fig. 1. The overbreak zone represents the zone beyond the minimum excavation line of the designed periphery from where rock blocks/slabs detach completely from the rock mass. It is a measure of difference in excavation between ‘as designed profile’ and ‘as excavated profile’. Overbreak zone is undesirable and leads to cost over-run due to extra excavation and backfilling, shotcrete, concrete or other material as per designed support system. Overbreak varies from 5% to 30% which incurs significant cost and increases cycle time of the tunnelling operation (Ramulu et al., 2009). The damaged zone is a zone around tunnel beyond overbreak zone. The irreversible changes in the rock mass properties take place in this zone due to presence of network of micro-cracks and fractures induced by the blasting excavation process. This zone is characterized by deterioration in mechanical and physical properties and increase in transmissivity properties (Saiang and Nordlund, 2009). The disturbed zone is a zone in the rock mass immediately beyond the damaged zone where changes in the rock mass properties are insignificant and reversible. This zone is dominated by changes in stresses and hydraulic permeability (Palmström and Singh, 2001). Overbreak as well as damaged zone has significant impact on the project cost, construction time, safety and performance of the underground structures. During construction of tunnels and caverns, damaged zone can adversely affect the stability of underground openings Enlarged extent of the damaged zone endangers safety of the front line workers as it may considerably reduce stand-up time of the rock mass. Functionality and post-construction performance of the structure will also be affected with enlarged extent of the damaged zone. The acceptable limit of damage to the rock mass varies with the importance and requirement of the excavation in different industries (Olsson and Ouchterlony, 2003; Mandal et al., 2005). During construction of a high level nuclear waste disposal system, even the 150

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Predominant rock mass encountered, and range of the data and Q values used in the study has been givenin Table 1. The tensile strength, Pwave velocity and Young’s modulus of rock mass encountered at the experimental sites have been presented in Table 2. Two types of rock masses namely thinly bedded phyllitic quartzite (PQT) and massive phyllitic quartzite (PQM) have been encountered in PSP project sites. Q-values of these two types of rock masses varied between 0.03 and 1.68 (Table 1) showing extremely poor and fair qualities of rock masses. However, at Singoli-Bhatwari project site, and Tapovan-Vishnugaad project site rock mass encountered represented fair to good rock mass having Q-values in the range of 2.7–17.8 (Table 1). The P-wave velocities of the rock mass have been found to vary in the range of 3.0 km/s to more than 6.0 km/s (Table 2). Rock mass characterisation, blast vibration monitoring, overbreak assessment and estimation of damaged zone (as discussed in Section 3) was carried out during each blast operation. All the experimental blasts under observation were closely monitored. Efforts were made to carry out experimental investigation encompassing large range of Q values which are generally encountered in tunnelling. Thus, range of Q values covered in the present study varied from 0.03 (extremely poor rock mass) to 17.4 (good rock mass). The cross-sectional area of tunnel varied from 25.9 sq. m. to 67.8 sq. m.

adequately address the in-situ conditions (Reference). Maerz et al. (1996) and Scoble et al. (1997) have emphasized that damage assessment method should be accurate, simple in procedure and should have a possibility to function under a large range of conditions. This will help the site personnel in making damage assessment a routine practice and timely implementing the control measures for reducing damage. Damage in the rock mass is influenced by several factors such as blasting practices, rock mass properties, geological conditions and also by stress environment around an opening. The blast induced damage happens with the blast, whereas the stress induced damagetakes place with some time with the interaction of rock mass and stresses in the process of attaining a fresh equilibrium state. Thus, the impacts of blasting parameters seem to be significant in the immediate rock mass damage around an underground opening. The stress induced damage may be spread to a greater extent depending upon ground conditions (Hoek and Karzulovic, 2000; Singh and Xavier, 2005). The combined effect of blast and stresses has not been evaluated in this work separately. Stress environment and rock mass properties cannot be altered after finalization of the project parameters but the blasting practices can be suitably refined to reduce damage (Backstrom, 2008). Therefore, it is important to study the inter-relationship between extent of damage and parameters of blast design in order to evaluate and limit the damage. Researchers have emphasized on the need of quick, simple, inexpensive and robust system for blast damage assessment (Mandal et al., 2008). In the light of above observations, efforts have been made in this study to develop empirical relations for estimation of extent of damaged zone based on the data obtained from the field investigations at five under construction hydropower project tunnels in Himalayan region of north India. It is appropriate to mention here that the commonly used explosives in these tunnels are small diameter cartridge slurry and emulsion explosives. The field-oriented research work is planned to develop simpler empirical approach to predict blast induced damage. Therefore, it was important to select parameters which are easily available to practicing site engineers without any laboratory tests. This empirical correlation will also help to integrate support design process with routine blasting operation with due consideration to blast induced damage.

2.2. Selection of parameters In underground construction using drill and blast method, rock mass damage around an underground opening primarily depends upon rock mass conditions, stresses and blasting parameters. In this study, efforts have been made to include all the influencing parameters to develop empirical correlation. Parameters related to blast design such as hole depth, spacing and burden of each hole, delay timing and firing sequence are collected from each experimental blasting operation. The data for drilling patterns including spacing and burden, (with emphasis on holes in perimeter and penultimate row and hole depth) was collected during drilling operations. Parameters related to explosive consumption in a hole as well as in total round, initiation system and firing sequence, maximum charge per delay were recorded during charging of holes in a blast round. Record of pull in each round was obtained after surveying of tunnel profile and advancement. Parameters such as total charge used in a blast round (T), maximum charge per delay (W), Pull (l), hole depth (d) were directly available. Other parameters such as advancement factor, confinement factor and perimeter charge factor were calculated from the recorded observation for each round of blast. In the proposed empirical correlation, specific charge (q), maximum charge per delay (W), perimeter charge factor (qp), advancement factor (Af) and confinement factor (Cf) have been used to represent the blasting operation in underground excavation. The parameters used are described below.

2. Field investigations and selection of parameters 2.1. Field investigations Field experiments have been carried out to gain insight of these influencing parameters at five tunnel construction sites. The sites are integral parts of three major hydro power projects located in Himalya, Uttarakhand state, India. These sites are as follows:

• Access tunnels AA10R and AA7 from pump storage plant (PSP) project of THDC India Limited at Tehri, • Head race tunnel (HRT) of Singoli-Bhatwari hydroelectric power project (SBHEP) at Rudraprayag, • Head race tunnel (HRT) and bypass tunnel (BPT) of Tapovan-

• Specific Charge (q) (kg/m ): Specific charge is defined as ratio of 3



Visnhnugaad hydroelectric power project (TVHEP) at Tapovan.

The data was collected from 113 different test blasts undertaken at five tunnel construction sites as shown in Fig. 2. The test sites have been selected based on suitability of rock mass conditions for blasting experiments, relatively lesser overburden so that stress environment around an opening has negligible influence on the damage distance and the damage is predominantly governed by the blasting operation. The minimum and maximum overburden in the experimnetal tunnels were 56 m and 318 m respectively. There was lesser variation in the topography of the experimental sites and therefore influence of topography was not considered in the study.

• • • 151

total quantity of explosive used and rock volume broken. It is expressed in kg/m3. Maximum Charge per Delay (W) (kg): It is maximum quantity of explosive fired in a delay series. This is obtained from record of delay distribution in a blast round and charging pattern in each hole. Perimeter Charge Factor (qp) (kg/m3): Similar to specific charge, perimeter powder factor is the quantity of explosive used in perimeter holes and the volume of rock corresponding to burden of the contour holes. Advancement Factor (Af): It is ratio of pull (l) and hole depth (d) in a blast round. Confinement Factor (Cf): It is ratio of hole depth (d) and cross-sectional area of tunnel (a).

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Singoli-Bhatwari HEP (99 MW) Range of Q Rating: 3 to 10

TVHEP, Joshimath, (520 MW) Range of Q Rating: 2.6 to 17.8

THDC PSP (1000 MW) Range of Q Rating: 0.03 to 1.67

Fig. 2. Location of Field Experimentation Sites.

Table 1 Details of experimental project sites and range of data. Source: detailed project report of respective projects (L & T- SBHEP, 2007; NTPC Ltd, 2006, 2010; THDC Pump Storage Plant Project Report, 2012). Sl. No

Project Site

Tunnel

Rock Type

Number of Test Data Set

Range of Q

1 2 3 4 5

Pump Storage Plant Project (PSP), Tehri

AA7R AA10R HRT HRT BPT

Thinly bedded Phyllitic Quartzite (PQT) Massive Phyllitic Quartzite (PQM) Quartz Biotite Schist Augen Gneiss Quartzite

27 30 20 24 12

0.8–1.1 0.03–1.68 2.7–11.1 3.6–4.3 6.8–17.8

Singoli- Bhatwari Hydro Power Project (SBHEP), Rudraprayag Tapovan Vishnugaad Hydro Power Project (TVHEP), Tapovan

Table 2 Geotechnical properties of rock in experimental tunnels. Source: detailed project report of respective projects (L & T- SBHEP, 2007; NTPC Ltd, 2006, 2010; THDC Pump Storage Plant Project Report, 2012) Sl. No

Experimental Tunnel Site

1

HRT SBHEP

2 3 4 5

HRT TVHEP BPT TVHEP AA7 PSP AA10R PSP

σt MPa

Vp m/s

E MPa

Vcr mm/s

−1.20

6.71

3267

12600

1739.8

−1.598

8.7

5400

27900

1683.8

−1.334

12.4

6200

55500

1754.5

−1.16

4.3

5400

10500

221.65

−1.03

7.2

6000

12700

340.15

Vibration Attenuation Equations

( ) = 1390.5 ( ) = 2079.1 ( ) = 441.9 ( ) = 576.2 ( )

Vppv = 1825.1 Vppv

Vppv Vppv

Vppv

R

W 0.33 R

W 0.33

R

W 0.33 R

W 0.33

R

W 0.33

Notations: σt: Tensile strength, Vp: P-wave velocity, E: Young Modulus; Vcr: Critical peak particle velocity; SBHEP: Singoli-Bhatwari Hydroelectric Project; TVHEP: Tapovan Vishnugaad Hydroelectric Project, Tapovan; PSP: Pump Storage Plant Project, Tehri.

(Maerz et al., 1996; Ibarra et al., 1996). The perimeter charge factor depends upon advancement & confinement factors and maximum charge per delay. Therefore, perimeter charge factor along with maximum charge per delay, advancement and confinement factors has been included in the proposed empirical correlation. Underground blasting operation is rela7tively difficult blasting operation as compared to surface blasting due to lack of free face. The

The above parameters are recorded in all the blasting operations as routine practice and are easily available to the site engineers. In the proposed empirical correlation, hole depth, confinement factor and advancement factor represent blast design parameters whereas maximum charge per delay, specific charge and perimeter charge factor represent explosive parameters. Effect of perimeter charge factor on overbreak and underbreak in underground excavation is significant 152

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Where, Vppv = Peak particle velocity, mm/s, K , β = Site constant (function of characteristics of propagating media), R = Distance of measurement, md andW = Maximum charge per delay, kg. The constant K and β have been determined using least square regression analysis for all the sites. Pal Roy (2005) suggested that in order to improve the confidence level in vibration prediction model, the regression line is moved up parallel. If the line is moved up by one standard error, then 84% of the actual values would be below the line. Similarly, moving up the line 1.65 times the standard error, the confidence level would enhance to 95%. In the present study, the attenuation characteristic is derived at 95% confidence interval and same is used for prediction of the blast vibration in all the five sites. In this study, attenuation characteristic is derived at 95% confidence interval and same is used for prediction of the blast vibration at all the five sites. The attenuation characteristics of the vibration obtained at five sites are presented in Table 2. Blast induced rock mass damage is a result of the induced dynamic stress during detonation. For an elastic medium, induced dynamic strain can be calculated as a function of peak particle velocity (Vppv) and longitudinal wave velocity (????). Therefore, blast induced damage arising in the rock mass is widely correlated with peak particle velocity of blast induced vibration. In this study, extent of damage to the surrounding rock mass is calculated using critical peak particle velocity (vcr) (Holmberg and Persson, 1978; Singh, 1993; Kwon et al., 2009). Critical peak particle velocity for each of experimental predominant rock mass is obtained using Eq. (3).

relative ease of blasting operation increase with tunnel size (a). The specific charge in a blast round is also function of tunnel size. Specific charge reduces with increase in tunnel size (Olofsson, 1988; Chakraborty et al., 1994). The extent of rock mass damage is significantly influenced by specific charge and confinement factor. Both the terms measure efficiency of blasting operation. A well utilized explosive energy in blasting round will produces lesser damage which is measured by advancement factor. The parameters such as specific charge, perimeter charge factor and maximum charge per delay are mutually exclusive and hence all the three parameters incorporated in the empirical correlation. Rock mass characterisation has been carried out using rock mass quality system, Q, developed by Barton et al. (1974). Parameters of Qsystems are rock quality designation (RQD), joint set number (Jn), joint roughness number (Jr), joint alteration number (Ja), joint water condition (Jw), and stress reduction factor (SRF). All these six parameters of Q system are given individual rating and the numerical value of Q is obtained using Eq. (1). In Eq. (1), the term (RQD/ Jn) is a measure of block size, (Jr/Ja) represents inter block shear strength and the term (Jw/ SRF) is a measure of active stresses in the rock mass. SRF is also a measure of rock burst, squeezing or swelling conditions Therefore, SRF is regarded as total stress factor (Singh and Goel, 2011). The value of Q varies from 0.001 to 1000.

RQD ⎞ ⎛ Jr ⎞ Jw ⎞ ⎛ Q=⎛ ⎝ Jn ⎠ ⎝ Ja ⎠ ⎝ SRF ⎠ ⎜

⎟⎜



(1)

Q-system of rock mass characterisation has been recommended specifically for tunnels and caverns with an arched roof (Singh and Goel, 2011). It is observed that Q-system is preferred method of rock mass classification for tunnels and caverns and it has been used to design more than 1200 structures world-wide (Kumar, 2002). A large number of field and design engineers as well as geologists are using Q-system for support design and engineering analysis of rock mass structures. Prevailing stress environment influences damage to the surrounding rock mass. In Q-system, stress reduction factor (SRF) is one of the parameters which accounts for active stresses during construction of an underground opening. Therefore, Q-system has been selected for rock mass characterisation in the present study. Moreover, number of correlations using Q parameter are available which can be used for obtaining other engineering rock mass properties used in analysis of underground excavation and design of supports.

Vcr =

During field investigation in each tunnel, vibration monitoring has been carried out at all the sites for determination of attenuation characteristics of blast induced ground vibration. Blast induced ground vibrations were measured using three different engineering seismographs namely MinimatePlus (MMP), Minimate Blaster (MMB) and Minimate (MM), (manufactured by Instantel Canada). Monitoring of the blast induced ground vibration was carried out as per the guidelines given in IS: 14881 (2001), ISRM suggested method (ISRM, 1992) and standard operating pProcedures recommended in Instantel Minimate User Manual (Instantel, 2009). Langefors and Kihlstrom (1963) proposed square root scaled distance. Ambraseys and Hendron (1968) further modified and suggested a model for prediction of blast induced ground vibration attenuation for spherical charge geometry using cube root scaled distance. Recently, Liang et al. (2011) conducted a comparative study of different scaled distances and recommended use of “cube root scaled distance law” for vibration study in tunnels as suggested by Ambraseys and Hendron (1968). The suggested blast vibration prediction model is as given in Eq. (2).

4. Analysis of the data The data obtained was analysed to gain insight into the influence of the major blast design parameters to the damage induced by the blasting operation. Analysis of data reveals that maximum charge per delay, perimeter charge factor, rock mass quality and advancement and conferment factors significantly influence damage to the rock mass around an opening. The contributions of the major parameters are discussed in the following sections. 4.1. Influence of rock mass quality (Q) Fig. 3 shows the variation of average damage distance with rock mass quality, Q. In general, the damage distance decreases with the increase in Q-values. The average damage distance for rock mass having Q-value less than 1 is greater than 5.0 m. It sharply reduces to approximately 3.0 m for rock mass with Q-value greater than 4. The impact of rock mass quality is significant in lower classes of the rock mass having Q-value less than 4. In higher classes of rock mass, impact of the rock mass quality remains fairly uniform and possibly other parameter

−β



(3)

Where, Vcr=Critical peak particle velocity, mm/s VP = P-wave velocity of rock, m/sσt = Tensile strength of rock, MPa, andE = Young’s modulus of rock, MPa. Geotechnical properties of the predominant rock mass encountered at experimental tunnel site are presented in Table 1. The critical peak particle velocity (Vcr) value was obtained using Eq. (3) for each experimental sites. The damage distance from a blast round can be back calculated using Eq. (2). Although the damage distance can be calculated using Eqs. (2) and (3), efforts have been made in this study to develop an empirical correlation based on readily available site parameters to directly estimate the damaged distance/zone. Following this method, damage distance for all the observed blasts has been obtained.

3. Estimation of damage distance (Dd)

R ⎞ Vppv = K ⎛ 3 ⎝ W⎠

VP σt E



(2) 153

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Avergae Damage Distance, (Dd ), m

6

sequential firing of the holes using adequate delay series. As the number of delay series reduces, maximum charge per delay increases and thus the blast induced damaged zone increases. Differential response of the rock mass for increasing values of the W is governed by the significant difference in the threshold values of damage. In case of poor rock mass, such as observed in phyllite rock in AA10R and AA7 tunnels of PSP project, the threshold value is less than 500 mm/s of peak particle velocity. Whereas in case of gneiss and quartzite rock mass encountered in other tunnels, the critical peak particle velocity of the rock which causes damage is greater than 1500 mm/s. Thus in lower class of rock mass, damaged zone is much larger as compared to the higher classes of rock mass for a given range of W. In tunnel blasting operation, firing sequence of the holes should be such that the maximum free face is created towards the bottom of tunnel section and hence maximum stoping can be done in favour of gravity (Chakraborty et al., 1994). Generally, maximum charge per delay is contributed by the bottom floor holes. Improper firing sequence increase confinement of blasting operation and maximum charge per delay which in turn increases extent of blast induced damage to the surrounding rock mass. In general, maximum stoping should be completed before firing the contour holes in tunnel blasting. This enables contour holes to create fracture using minimum explosive and hence least possible damage is inflicted to the remaining rock mass. All the holes at the perimeter of tunnel are blasted in same delay. Higher charge concentration in perimeter holes may also contribute to maximum charge per delay. Such blasting operation will have compounding impact on the blast damage zone and the resultant damage to the rock mass may be more than 4.0 m.

5 4 3 2 1 0 <1

1-4 4 - 10 Rock Mass Quality Index, Q

> 10

Fig. 3. Variation of Damage Distance (Dd) with Rock Mass Quality Index, Q.

of blast design plays pivotal role in defining damaged zone around an opening.

4.2. Influence of maximum charge per delay (W) In any underground blasting operation, progressive enlargement of the free face shall be achieved by designing the firing sequence of holes using different delay series. Proper distribution of the delay series ensures free face to each hole. The holes are fired in the direction of free face thus utilising the explosive energy in breaking and displacing the rock. The maximum charge per delay, W depends on the number of delay series used in a blast round. Improper delay distribution gives excessive burden and spacing to the holes which leads to generation of the blast induced ground vibration resulting into greater extent of damaged zone. Thus, maximum charge per delay, W, influences significantly the damage distance among other parameters. The variation of observed average damage distance with maximum charge per delay is shown in Fig. 4. In experimental blast rounds, range of W varied between 15 kg to 40 kg. Therefore, a range of 15–40 kg maximum charge per delay, W has been considered in the analysis. It may be noted from Fig. 4 that the damage distance, as expected, increases with increase in maximum charge per delay. In tunnel blasting, it is important to achieve progressive enlargement of tunnel (Chakraborty et al., 2004). This can be achieved by

4.3. Influence of blast design parameters Fig. 5 is a plot of damage distance with parameters related to blast design which include hole depth, advancement and confinement factors and maximum charge per delay. The parameters are normalized with rock mass quality index to make data comparable for different rock mass conditions and tunnel sizes. In this study, experimental investigations have been carried in a wide range of rock mass quality (0.03–17.8) and tunnel size (25.9–67.8 sq. m.). The blast design parameters such as hole depth, advancement rate, confinement factor depend primarily upon rock mass quality, tunnel size, etc. Therefore blast design parameters and rock mass quality are grouped together as they are mutually inclusive parameters. The exponent to blast design and

5

Damage Distance, (Dd ), m

4.5

4

3.5

3

2.5

2 < 15

25

30

35

40

Maximum Charge per Delay (W ), kg Fig. 4. Variation of damage distance (Dd) with maximum charge per delay (W).

Fig. 5. Variation of Damage Distance with Blast Design Parameters.

154

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rock mass parameters used in the expression for predicting damage distance is obtained by multiple linear regression analysis technique to get best fit curve. In general, damage distance increases with increase in advancement factor due to increase in confinement factor, total charge and maximum charge per delay. This can be attributed to the fact that in experimental blasts, higher advancement has been achieved in most cases due to increase in either hole depth or total charge used in blast round. Higher value of total charge in a blast round will increase advancement but will also cause more damage to the rock mass. It is equally important to achieve optimum pull in every blast round to utilise the explosive energy in useful productive work such as rock breakage and throw. Under-utilised explosive energy will increase damage to the rock mass due to increase vibration intensity (Berta, 1990). An optimum advancement rate achieved through optimized blast design parameters will reduce damage in the surrounding rock mass. The reason behind the decrease in damage may be due to the fact that the increase in advancement factor also leads to optimum utilization of the explosive energy which otherwise would have converted in blast induced ground vibration. Fig. 6. Damage Distance versus Explosive Energy Parameters for Q < 1.67.

4.4. Influence of explosive parameters In blasting operation, explosive energy appears in three forms, maximum charge per delay (W), specific charge (q) and perimeter charge factor (qp). Whereas maximum charge per delay is influenced by the initiation and firing sequence of the blast round, perimeter charge factor and specific charge are dependent on advancement of a blast round. Maximum charge per delay and specific charge are measures of explosive energy in the blast round whereas the perimeter charge factor is parameter introduced to measure the damage created by the explosive energy in the contour hole. For a fixed amount of explosive energy, greater advancement rate reduces perimeter charge factor as well as specific charge. All the three explosive energy parameters W, qp and q are grouped as factor Z as given in Eq. (4). The vibration intensity in terms of peak particle velocity is proportional to maximum charge per delay (Langefors and Kihlstrom, 1963; Devine, 1966 and Dowding, 1985). The basis of such formulation is commonly used square root scaled distance concept incorporating all the collaborating parameters of blast design

Z = qp0.15 W + q

(4)

Where, qp = Perimeter charge factor, kg/m3, W = Maximum charge per delay, kg, and Q = Specific charge, kg/m3. In case of tunnels AA10 R and AA7, rock mass was of lower quality (Q < 1.67) whereas in other three experimental tunnels Q values were higher than 1.67. Data for these two cases are plotted separately due to difference in rock mass classes and also in blasting practices. In experimental tunnels AA7 and AA10 R rock masses are of very poor category (0.03–1.67), cross-sectional area are higher (55.7–67.8 m2) and hole depth is less than 2.5 m. In other tunnels, rock masses are of better quality, tunnel cross-sectional area is less than 40 m2 and hole depth were greater than 3.0 m. As shown in Figs. 6 and 7, factor Z (Eq. (4)) is directly proportional to the damage distance. In both the cases a correlation coefficient of 0.83 is obtained. In a blasting operation in underground excavation, contour holes are fired in the last delay series. All the holes are assigned same delay and spacing in these holes are lesser than the burden of the contour holes. Such firing arrangement creates a fracture line along the final excavation line. Such arrangements of delay for periphery holes, although these holes are lightly charged, contribute as maximum charge

Fig. 7. Damage Distance versus Explosive Energy Parameters for Q > 1.67.

per delay due to large number of holes fired in same delay time on several instances. These conditions compound the effect of maximum charge per delay with perimeter powder factor. Conventionally, contour holes are given last delay series so that the maximum rock breaking is done in favor of gravity. Even in smooth wall blasting technique, contour holes are fired in same delay. Due to large number of holes in periphery, although lightly charged, their contribution is comparable to that delay series, having maximum charge per delay. In such circumstances, the effect of the perimeter charge factor increases and resultant damage distance is significantly higher. Although the periphery holes provide a line of fracture along final line of excavation, due to it initiation in the last delay series, the effect of W is not restricted by the lightly charged contour holes. Therefore, the damage distance is enhanced by the perimeter charge factor. Differential response of the rock mass for increasing values of the W is governed by the significant difference in the threshold values of damage. In case of poor rock mass, such as observed in phyllite rock in AA10R and AA7 tunnels of PSP project, the threshold value is less than 155

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are incorporated in the form of advancement factor and confinement factors. Optimum advancement rate reduces damages, whereas, higher confinement factor reduces ease of blasting operation and requires additional explosive energy to achieve equivalent advancement rate as compared to blast with less confinement rate. Eq. (6) can be further simplified as Eqs. (7) and (8)

⎡ qp0.15 W + q ⎛ da ⎞ Dd = 0.96 ⎢ ⎜l⎟ Q 0.33 ⎢ ⎝d⎠ ⎣

0.15

⎤ ⎥−1.28 ⎥ ⎦

0.15 W + q Cf 0.15⎤ ⎡ qp ⎛ ⎞ Dd = 0.96 ⎢ ⎜ ⎟ ⎥−1.28 0.33 Q ⎝ Af ⎠ ⎦ ⎣

(7)

(8)

Dd ⩾ 0 Confinement factor, Cf =

d a

l

Advancement factor, Af = d Where, d = Depth of drill hole, m, a = Tunnel cross-sectional area, m2, and Equation (8) can be used for prediction of the damage distance induced by blasting in underground excavation. The proposed empirical correlation, as given in Eq. (8), encompasses parameters related to rock mass quality (Q), explosives energy (W, qp and q), blast design (d and Cf), tunnel size (a) and also results of the blasting operation (Af) which all influences damage to the rock mass in a tunnel blasting operation. The proposed correlations are based on readily available site parameters which may be helpful to the practicing engineers and geologists in the optimization of the support design of rock mass. Eq. (8) gives impression that energy parameters are used repetitively in the recommend correlation. It may be noted that all these three parameters are mutually exclusive. In a same blast design, values of W, qp and q can be altered without changing other parameters. In blast round, having same drill hole depth and total charge, arrangement of firing sequence will change the values of W. Pull of the blast from such changed configuration will alter values of qp as well as q. Inclusion of these three parameters will therefore be able to assess their effect on blast induced damage distance. Endeavour of an excavation engineer is to achieve maximum advancement in each round of blasting operation so that the production is maximum and progress is faster. It is suggested that the blast design parameter shall be optimized for achieving progress in each round rather than merely increasing explosive quantity. Achieving advancement through optimized blast pattern is also advantageous in reducing blast induced damage to the surrounding rock mass.

Fig. 8. Plot of Factor D and observed damage distance, Dd.

500 mm/s of peak particle velocity. Whereas in case of gneiss and quartzite rock mass encountered in other tunnels, the critical peak particle velocity of the rock which causes damage is greater than 1500 mm/s. Thus in lower class of rock mass, damaged zone is much larger as compared to the higher classes of rock mass for a given range of W. 5. Empirical correlation for prediction of damage distance (Dd) As discussed in the preceding section, damage distance is influenced by the blast design parameters and rock mass quality. Analysis of various parameters considering the general trend of the field data and also the damage mechanics was performed to develop an empirical correlation for prediction of Damage distance Dd using all the parameters. Exponent of blast design parameter in the expression are obtained by sensitivity test and using multiple linear regression analysis technique to represent the best fit curve. Fig. 8 shows plot of a factor D (Eq. (5)) and damage distance (Dd) obtained as discussed in Section 3. In Fig. 3 all the observed Dd values obtained using field investigation at five tunnel construction sites have been plotted and the equation of best fit line is obtained as Eq. (6).

D=

qp0.15 W + q Q 0.33

2 0.15

⎛d ⎞ ⎝ al ⎠ ⎜



(5)

0.15 W + q d 2 0.15⎤ ⎡ qp ⎛ ⎞ Dd = 0.96 ⎢ −1.28R2 = 0.88 0.33 Q al ⎠ ⎥ ⎝ ⎣ ⎦ ⎜

6. Validation of empirical correlation



(6)

Fattahi et al. (2014) carried out a study for selection of a suitable method for the assessment of excavation damage zone using fuzzy AHP in Aba Saleh Almahdi tunnel in Iran. They compared various methods of assessment of excavation damage zone for a tunnel in Iran using fuzzy analytical hierarchy process. They found the geophysical method and borehole core drilling and logging as reliable one amongst other techniques. The proposed empirical correlation has been validated using geophysical test on rock core samples obtained from ten locations from HRT of SBHEP project. Ultrasonic tests have been performed on the rock core samples following guidelines as per ISRM suggested method (ISRM, 1981). Percentage reduction of P-wave velocity with depth was computed from the ultrasonic test data. As per Liu et al. (2009) and Fu et al. (2014), the threshold of damage is defined as a 10% reduction in the P-wave velocity as compared to the P-wave velocity of the undisturbed rock mass and computed at each coring locations. It has been assumed that rock mass at maximum drill depth represents undisturbed

Dd ⩾ 0 Where, Dd = Damage distance, m, qp = Perimeter charge factor, kg/m3, W = Maximum charge per delay, kg, q = Specific charge, kg/m3, Q = Rock mass quality index (Barton’ Q-system) d = Hole depth, m, l = Pull, m, and a = Tunnel cross-sectional area, m2. The factor D (Eq. (5)) have been formulated on the basis of correlation as discussed in the preceding section and principle of the construction blasting operations. It may be noted that that the explosive energy parameters (W, q and qp) are proportional directly to damage distance. The rock mass quality index bears inverse relationship with extent of the blast induced rock mass damage as good quality rock mass can sustain higher level of the vibration. The blast design parameters 156

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4

this study. The combined effect of blast and stresses has not been evaluated in this work separately. Many researchers have recommended to neglect SRF during rock mass characterisation and to assess the effect of stress separately. Tolerance limit of the damage depends upon the purpose and service life of the underground structures. Therefore, further research is recommended to incorporate such aspect in the studies pertaining to damage assessment.

Damage Distance, m

Predictd Damage Distance Observed Damage Distance

3

2

Acknowledgements Authors express their gratitude to the project authorities of National Thermal Power Corporation Ltd, New Delhi, THDC India Ltd., Rishikesh, M/s L & T, Tapovan-Vishnugaad Hydropower Project. Team of engineers and geologists from L & T Singoli-Bhatwari Hydropower Project have also provided great help in field investigations. Their contribution is also thankfully acknowledged.

1

0 424.8

22.0

25.0

350.4 353.3 359.0 431.4 472.1 474.7 478.0

Blast Location (RD, m)

References

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rock mass. The corresponding average P-wave velocity, therefore, is said to be the velocity of the rock mass before blasting operation. The distance up to which there is 10% reduction in the P-wave velocity is designated as observed damage distance. Similar approach has been adopted for estimation of the observed damage distance for all the ten locations. The damage distance obtained from the ultrasonic test data were in close agreement with the predicted damage distance obtained using Eq. (7). A bar chart to evaluate the relative difference between predicted using proposed correlation (Eq. (7)) and observed damage distance have been presented in Fig. 9. It can be seen in Fig. 9 that the values of the predicted damage distance and observed damage distance are in close agreement. 7. Conclusions A comprehensive field investigation have been carried out at five tunnel construction sites to evolve empirical correlation for estimation of damage distance using readily available site parameter. Observations of 113 blasting experiment have been taken in different rock mass from extremely poor to good rock mass class. An empirical correlation has been suggested using Specific Charge (q), Maximum Charge per Delay (W), Perimeter Charge Factor (qp), Advancement Factor (Af) and Confinement Factor (Cf) and rock mass quality index Q. Proposed correlations will also enable practicing engineers in carrying out damage audit on the regular basis with optimal instrumentation which in turn can be integrated with the support design process. Moreover, the rock mass having same intact rock constituents but different fracture configuration and stress state will develop different extent of damage subjected to given vibration intensity. The present study attempts to include the effect of pre-existing fractures pattern and stress state by imbibing the rock mass quality index Q in the analysis. Using proposed correlation, the extent of damaged zone may be computed for known values of Q and blast design parameters. The blasting operation reduces the rock mass quality which is reflected in reduction in Q -rating. The reduction in Q-rating and estimated extent of the damaged zone shall be integrated during support design. Therefore, support design engineers shall consider deterioration in rock mass quality with respect to type of blasting, for an effective control of overbreak and enhance safety and stability of underground structures. In the present study, a wide range of Q values (0.03 to 17.8) has been considered. The impact of stresses on extent of damage around an opening may be significant in deeep seated tunnel. The stress is accounted for during rock mass characterisation as SRF as a part of Q in 157

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