TBIpdf

Development of rational models for tunnel blast

prediction based on a parametric study

A. K. CHAKRABORTY, A. K. RAINA, M. RAMULU, P. B. CHOUDHURY,A. HALDAR, P. SAHOO and C. BANDOPADHYAYCentral Mining Research Institute, Regional Centre, 3rd Floor MECL Complex,Seminary Hills, Nagpur-440 006, India. e-mail: [email protected]

(Received 25 February 2003; revised and accepted 25 June 2003)

Abstract. The empirical models available for prediction of the tunnel blast results like pull

ratio, specific charge, specific drilling and overbreak have some inherent shortcoming inabsence of any parametric study at the backdrop. Hence, the models use different constitutingparameters and provide values which differ widely. After a thorough review of literature and

field investigations in the drivages of mines and tunnels some parameters were identified.Those parameters were subjected to Multiple Linear Regression analyses to filter out the mostinfluencing ones which represent the rockmass properties, the tunnel configurations and theblast designs. A parameter called Tunnel Blasting Index (TBI) was conceptualized and was

expressed in terms of those most influencing parameters. All the blast results observed duringthe filed investigations could be well related to a single index TBI. Some adjustments onaccount of shape of the tunnel and joint orientations, which were not addressed in the avail-

able models, are suggested in the developed models.

Key words. joint orientation adjustments, most influencing parameters, predictive models,Tunnel Blasting, Tunnel Blasting Index (TBI).

1. Introduction

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 and overbreak or underbreak.

The underbreak is usually expressed in the field in terms of negative overbreak.

The blast induced overbreak or underbreak is measured radially and expressed

in metre. Those are also occasionally estimated volumetrically in m3 of in situ

rock mass over or under-broken and expressed in percent of the designed

volume. However, in most of the projects in India, the permissible limit of over-

break has been defined in terms of width and height of tunnel. The Swiss Society

Geotechnical and Geological Engineering 22: 477–496, 2004. 477# 2004 Kluwer Academic Publishers. Printed in the Netherlands.

of Engineers and Architects defines the permissible overbreak limit as 0:07pA,

where A is the tunnel area or 0.4m whichever is less (Innaurato et al., 1998).

All the above results jointly contribute to the safety, rate of progress and economy

of the tunnel. Hence, it would be misleading to measure the tunnel blast efficiency in

terms of one or two of the above parameters. The degree of fragmentation and the

muck profile are also important indicators of blast performance and may need to be

addressed in several cases, particularly in open pit mining operations. However, as

these two results affect the mucking operation and generally do not pose severe pro-

blems to the practising engineers in their day-to-day tunnel blasting activities, these

are not dealt in details in the present paper.

The above results, are dependent on the rock mass properties which may be

termed as the non-controllable parameters (N), the tunnel configurations which

may be called as the semi-controllable parameters (S) and the blast design para-

meters which are known as the controllable parameters (C). All the parameters used

for development of available models leading to blast design or predictions can be

classified in these three categories as is evident in Table 1.

The following aspects are prominently noted in Table 1.

(1) The non-controllable parameters have been adequately considered in most of the

available predictive models, but the semi-controllable or the controllable para-

meters are not amply represented.

(2) The degrees of influence of the parameters considered in those predictive models

are not known. It is also known if those models considered the most important

parameters. Further, the interrelations among the influencing parameters are

also not revealed.

(3) No index is available to relate all the blast results. Hence, it is difficult to assess

the overall change in tunnel blast results if one of the parameters constituting

any model is varied. For example, it may be possible to assess the change in

specific drilling or specific charge if blast hole diameter is changed but its effect

on pull or overbreak can not be estimated as the predictive models for those do

not consider the blast hole diameter, though the linear charge density increases

with the increase in hole diameter.

It, therefore, appears imperative to conduct a parametric study to define the most

influencing parameters and subsequently develop the predictive models for all the

blast results with the help of those influencing parameters. This also enables to define

and weight all the blast results on the same platform. The present paper describes the

predictive models developed by the authors to fulfil the objective, from concept to

the actual development through parametric studies.

2. Development concept

In line with the Blast Damage Index ðDibÞ developed by Yu and Vongpaisal (1996),

the tunnel blast can be defined as a function of blast induced stress and tunnel rock

478 A. K. CHAKRABORTY ET AL.

Table 1 Parameters considered for prediction of tunnel blast results

Sl. no. Blast result

Parameters considered in

predictive models Model developed by

1 Specific charge

and specific drilling

Tunnel area (S)

Drilling error (S)

Langefors & Kihlstrom (1973)

Tunnel area (S) Olofsson (1988)

Tunnel area (S) Pokrovsky (1980)

Protodyakonov Index (N)

Rock structure (N)

Relative weight strength

of explosive (C)

Explosive (charge) diameter (C)

Tunnel area (S) Hagan (1992) and

Du Pont (1977)

Blast hole diameter (C)

Density of rock (N) Ghose (1988)

Protodyakonov Index (N)

Joint spacing (N)

Joint orientation (N)

Rock Mass Description (N) Lilly (1986)

Joint spacing (N)

Joint orientation (N)

Specific gravity of rock (N)

Hardness (N)

Rock Mass Quality (Q)(N) Chakraborty (1996, 1998)

Strength Rating (N)

Number of contact surfaces (N)

Hole length (C)

2 Pull Sonic velocity in rock (N) Bergh-Christensen and

Selmer-Olsen (1970)

Specific gravity of rock (N)

Number of joints in a round (N)

Uniaxial tensile strength (N)

Joint orientation (N) Johansen (1998)

3 Rock mass

damage and

overbreak

Peak particle velocity (N+C) Langefers & Kihlstrom (1973)

and Holmberg & Persson

(1978, 80)

Vector sum of peak particle

velocities in three orthogonal

(N+C) directions

Yu & Vongsaipal (1996)

Rock density (N)

P-wave velocity (N)

Dynamic tensile strength of rock (N)

Site Quality Constant (S)

Young’s Modulus (N) Mckenzie (1994)

Uniaxial tensile strength (N)

P-wave velocity (N)

Joint orientation (N) Johansen (1998)

Rock Mass Strength (N) Innaurato et al. (1998)

(N-Non-controllable parameter, S-Semi-controllable parameter, and C-Controllable parameter).

DEVELOPMENT OF RATIONAL MODELS 479

mass resistance to fragmentation, which jointly indicate the tunnel blast environ-

ment. Hence, tunnel blast results can be expressed as a function of rock mass resis-

tance to fragmentation and blast induced stress as shown in Equation 1.

Tunnel blast results

¼ f (tunnel rock mass resistance to fragmentation, blast induced stress)

ð1Þ

As tunnel blasting is done under confined conditions, the tunnel rock mass resistance

is controlled by not only the rock mass properties but also by confinement. It was

seen that DuPont (1977), Pokrovsky (1980), Langefors and Kihlstrom (1973) and

Olofsson (1988) considered the inverse of the tunnel area to account for the tunnel

confinement while predicting the specific charge and the specific drilling (Table 1).

Thus, Equation 1 can be modified as:

Tunnel blast results

¼ f f(rock mass resistance to fragmentation, tunnel confinement),

blast induced stressg ð2Þ

An index called Tunnel Blast Index (TBI) is conceived to represent the tunnel blast

environment comprising blast induced stress, rock mass resistance to fragmentation

and tunnel confinement. Hence, the tunnel blast results can be expressed as function

of TBI as shown in Equation 3.

Tunnel blast results ¼ f (TBI) ð3Þ

If the blast induced stress, the rock mass resistance and the tunnel confinement can

be expressed in terms of factors defined by the respective influencing parameters, TBI

can be defined by Rock Mass Factor (RF) which is a function of the rock mass prop-

erties providing resistance against fragmentation, Tunnel Configuration Factor (TF)

which is a function of tunnel configuration parameters contributing to the confine-

ment and Blast Design Factor (BF) which is a function of blast design parameters

responsible for blast induced stress.

As discussed earlier, it would be essential to identify the most influencing para-

meters to define RF, TF and BF. This needs collection of detailed information during

tunnel blasting under different environments. Field investigations were carried out in

tunnels of different configurations under various rock mass conditions. This was

followed by an analytical process to filter the parameters which have the maximum

influence on the observed blast results. TBI has been defined in terms of RF, TF and

BF which includes those most influencing parameters.

3. Field investigations

In accordance with the above scheme, field investigations were conducted by the

authors in inclined drifts of a coal mine, development galleries of two metal mines

and a tunnel of a hydro-electric project. The sites are listed in Table 2 and the loca-

tions of the sites are shown in Figure 1.


3.1. METHODOLOGIES FOLLOWED IN FIELD INVESTIGATIONS

The following broad methodologies were adopted by the authors during the field

investigations:

1. All the investigated tunnel lengths were thoroughly inspected.

2. Initial tunnel lengths, which mostly consisted of weathered rock masses, were

excluded.

Table 2 Names of the investigated sites

Sl. no. Type of mine/project Name of mine/project Nature of excavation

A. Mining sector

1 Coal Tandsi Inclined drifts in rock

2 Metal

(Manganese)

Chikla Development roadways, raises

and winzes in host rock and

drivages in manganese ore body

Gumgaon

B. Civil sector

3 Hydro-electric

project

Koyna Hydro-

electric

Project (KHEP),

Stage IV

Link tunnel in rock

Figure 1 Locations of the sites selected for field investigations


The blasting operation was not standardised in the initial portion of the tunnels.

The locations where the rock masses were found abnormally different from those

in the rest of the tunnel were not considered for further analysis. The locations,

which were ignored in this manner at different sites, vary from 5 to 20 percent

of the total population.

3. Rock Mass Quality (Q) was determined for all types of formations available in

the investigated tunnel lengths using the following reaction:

Q ¼ ðRQD=JnÞ � ðJr=JnÞ � ðJw=SRFÞ ð4Þ

where,

RQD ¼ Rock Quality Designation,

Jn ¼ joint set number,

Jr ¼ joint roughness number,

Jn ¼ joint alteration number,

Jw ¼ joint water reduction factor, and

SRF ¼ Stress Reduction Factor.

RQD values used for obtaining Q were determined from volumetric joint count

using the following relation provided by Palmstrom (1975):

RQD ¼ 115� 3:3 Jv ð5Þ

where,

Jv ¼ volumetric joint count.

The Jv values were determined by adding the number of visible joints per metre

length of the exposed surfaces in all three directions. The walls and the roof of

an excavation were scanned for this purpose.

4. The investigated tunnel length was categorised into various zones based on the Q

value. The joint set which appeared most frequently and consistently was consid-

ered as the major joint set in that zone. The orientation of that joint was deter-

mined using a Brunton compass.

The procedures followed for joint orientation and spacing measurements during

field investigations are in conformation with the guidelines provided by Interna-

tional Society of Rock Mechanics (ISRM) (Brown, 1981).

5. Rock samples representing all zones were collected by the authors.

The samples were tested in the laboratories of Central Mining Research Institute,

Regional Centre, Nagpur and Visvesvaraya National Institute of Technology,

Nagpur for evaluation of various physico-mechanical properties which were

found to be influencing the blast results as detailed in Table 1.

6. Detailed information on on-going blasting practice and four blast results like the

pull, the specific charge, the specific drilling and the overbreak in various rounds

were collected by the authors.


Face advance in a round was measured at the face centre and the two sides of the

face. The average of these values was considered as the average advance per

round. The roof and the side overbreak after each round of blasting were mea-

sured at 5 to 7 locations both side-wise and height-wise. These values were aver-

aged to obtain the average overbreak. The exacavated in situ volume was

calculated by multiplying the post-blast cross-section and average face advance.

This was verified from the number of trips of loaded muck or the stock pile or

crusher data. The specific charge or the specific drilling were estimated from the

ratio of total explosive quantity or total drilling length in a round and the exaca-

vated in situ volume of rock.

7. The blast results of different rounds in a particular zone were averaged to deter-

mine the average blast results in that zone.

8. Trial blasts were conducted in these sites with modified blast design and the

results were monitored by the authors.

The various geo-mining variations covered during field investigations are listed in

Table 3 to provide an overall picture at a glance.

During the field investigations many parameters listed in Table 1 were found to

influence the blast results significantly. However, some other parameters were also

found to affect the results as described below.

The directional blast results like the pull and the overbreak or the underbreak at

the roof or the walls are substantially influenced by joint plane orientation. The pull

was affected most adversely if the joints were steep with strike parallel to the tunnel

axis. Further, such joint sets striking across to the tunnel axis proved to be the most

favourable condition to improve the pull. An opposite trend was found in case of

roof overbreak. The roof overbreak was less when joints were steep but the strike

was parallel to the tunnel axis and was more when such joints had strike across

the tunnel axis.

Further, the tunnel shape and the application of contour blasting influenced

the overbreak. Also the unevenness at the tunnel periphery was fount to be lar-

gely dependent on the spacing to burden ratio. Holmberg (1982) recommended

different spacing to burden ratio for different parts of the tunnel specially for

the contour holes to minimize the overbreak. Additionally, the specific drilling

was also considerably increased in case of contour blasting as the spacing to

burden ratio of drill holes along the contour ðmdcÞ is maintained less than

one in contrast to those maintained as one or more in the rest of the tunnel

section.

4. Identification of the most influencing parameters

Based on the field investigations and the literature review, a list of influencing para-

meters has been prepared by the authors in Table 4.

The degree of influence of the above parameters on the blast observed in the field

investigations is determined by multiple linear regression analysis.


Table3Geo-miningconditionsobservedduringfieldinvestigations

C.Site

Sl.no.

Parameter

Link

Chikla

Tandsi

Gumgaon

A.Rockmassproperties

1Typeofrock

Basaltieflowof

compact&

amygdoloidal

basaltand

volcanic

breccin

Mn-oreandfootwall

containing

manganiferousquartz

andmuscoviteschist

Sandstone

Mn-orebodyand

footwallrock

masscontaining

quartzmuscovite

schist

2Q

5.378–64.48

0.75–32.05

0.31–18.66

0.21–1.85

3UCS,MPa

21.02–91.136

18–180

18.9–32.4

60–162

4Density,1/m3

2.37–2.93

2.5–3.9

1.9–2.35

2.82–3.97

5RQD

35.45–87.27

40.75–91.4

36–82

23–47.4

6P-wave

velocity,km/s

2.487–5.816

2.913–8.117

1.9–2.9

4.582–7.694

7Majorjointsetorientation

Dip()

60–90

60–90

0–30

30–60

Strikeangle

withrepectto

tunnelaxis()

0–30

60–90

60–90

30–60and60–90

8Mixedface

Brecciaþ

Amygdoloidal

basaltþ

Compactbasalt

Nil

Nil

Nil


B.Tunnelconfiguration

9Size,m2

36

5.04

15(with

shotcrete

support)and

17.66withsteel

support

5.04

10

Shape

Arch

Rectangular

D-shaped

Rectangular

11

Inclination

Nearlyhorizontal

Nearlyhorizontal

Inclination1:4.66

Nearlyhorizontal

C.Operatingtool

12

Drillingmachine

Manualjack

hammer

Manualjack

hammer

Hydraulicjumbo

Manualjackhammer

D.Blastdesign

13

Typeofcut

Convergent

Convergent

Parallelin

shotcrete

supportedzone

andconvergent

insteelsupportedzone

Convergent

14

Methodof

perimeterblasting

Conventional

Conventional

Conventional

lateron

switchedto

smoothblasting

Conventional


Multiple linear regression analysis (MLR) is performed to determine the combined

effect of a group of independent variables upon a dependent variable. The method

may be used to assign relative importance to the independent and may be interrela-

ted variables by sequentially including or excluding the one having largest partial

correlation (Gupta and Kapoor, 1999 and SPSS Inc., 1993).

The parameters considered for MLR are having different units and their ranges

vary widely. It is therefore considered that all these parameters should be normalised

using the following relation:

Xn ¼X

Xmax � Xminð6Þ

where, Xn is the normalised value, X is the original value and Xmax and Xmin are the

maximum and minimum values of that particular parameter in the population.

By normalising the variables and recasting them in dimensionless units, the arbi-

trary effects of similarities between the objects are removed (Flood and Kartam,

1994; Sayed and Abdewahab, 1998; Leu et al., 1998).

The gradual improvement in correlation between the specific charge and the spe-

cific drilling with the twelve input parameters obtained through MLR are presented

in Figures 2 and 3 respectively. The indices include the independent variables repre-

senting the X-axis in Figures 2 and 3.

It can be seen in the Figures 2 and 3 that, the index of correlation ðR2Þ improves

with the addition of independent variables. But the improvement is not all significant

Table 4 List of influencing parameters

Properties

Sl. no. Rock mass properties (N)

1 Density of rock mass

2 Rock strength–represented by Strength Rating (SR)

3 P-wave velocity

4 Barton’s Rock Mass Quality (Q)

5 Number of contact surfaces in multiple geological mixed rock face condition

6 Orientation of major joint set with respect to tunnel face

Tunnel configuration (S)

7 Area

8 Shape factor – the ratio between tunnel width to diameter of the curvature at roof

9 Tunnel inclination with respect to upward vertical direction

Blast design (C)

10 Type of cut

11 The deviation factor defined in terms of the ratio between drill hole length and diameter

12 Number of additional or baby cut holes blasted before the main cut holes

13 Charge per hole

14 Coupling ratio of explosive to blast hole diameter

15 Conventional or contour blast design expressed in terms of spacing to burden

ratio and coupling ratio in the contour holes


beyond the first six parameters, when the improvement occurs up to 1 percent. In

view of this fact the following seven parameters, which include all these first six para-

meters, are finally considered as most influencing. Those seven parameters are clas-

sified in three groups like, (i) rock mass parameters, (ii) tunnel configuration

parameters and (iii) blast design parameters as shown below:

4.1. (A) ROCK MASS PARAMETERS

1) P-wave velocity (cp, expressed in km/s)

2) Number of contact surfaces in multiple geological mixed face condition (n)

3) RQD (RQD)

4.2. (B) TUNNEL CONFIGURATION PARAMETERS

4) Area (A, m2)

5) Inclination ðbiÞ with respect to vertical upward direction (expressed in radian, r)

Figure 2 Results of multiple linear regression analysis for specific charge

Figure 3 Results of multiple linear regression analysis for specific drilling


4.3. (C) BLAST DESIGN PARAMETERS

6) Cut hole angle i.e., the angle made by cut holes with the face, expressed in cotan-

gent ðCaÞ

7) Coupling ratio between explosive and blast hole diameter ðRcÞ

The Rock Mass Factor (RF), the Tunnel Configuration Factor (TF) and the Blast

Design Factor (BF) are worked out from the above selected parameters, as per the

concept of the Tunnel Blasting Index discussed earlier.

Accordingly, the Tunnel Blasting Index (TBI) is defined as:

TBI ¼Rock Mass factor (RF)

Tunnel Configuration Factor (TF)� Blast Design Factor (BF)ð7Þ

where,

RF ¼ cp þ nþ ðRQD=10Þ; ð7AÞ

TF ¼ A� r; and ð7BÞ

BF ¼ Ca þ Rc: ð7CÞ

The P-wave velocity (cp) varied between 1000 to 8000m/s. Its range is exceptionally

wide in comparison to other parameters in TBI. Therefore, it is converted to km/s

unit to keep it in harmony with the other six parameters and to restrict the values

of TBI within reasonable limit. Among others, the range of RQD is also quite broad.

One tenth of RQD values are therefore considered in RF for the similar reason.

5. Development of models

In accordance of Equation 3, different blast results observed by the authors during

field investigations have been correlated below with TBI determined for the respec-

tive zones of the tunnels. Consequently new models for blast results prediction are

developed.

The relations between the observed specific charge (q, kg/m3) and the specific dril-

ling (bs, m/m3) with TBI are shown in Figures 4 and 5 respectively. Both the relations

have index of determination (R2) more than 0.9.

The specific charge predictive model (Equation 8) is developed on the basis of the

relation shown in Figure 4.

q ¼ 1.1+0.24 TBI, kg=m3 ð8Þ

As discussed earlier, a huge amount of additional drilling is essential in contour

blasting where the required explosive quantity is distributed in the closely spaced

holes along the perimeter. The spacing to burden ratio in the contour holes (mdc)

is included in specific drilling prediction to account for the additional drilling


required in contour blasting, if any. Further, the shape factor (sh) has also been

considered in view of the fact that additional drilling required at the perimeter in

an arch shaped roof than in a flat roof. Shape factor is defined as the ratio of the

tunnel width to the diameter of roof curvature. The ratio is 1 for a perfect D-shaped

Figure 4 Specific charge vs. TBI

Figure 5 Adjusted specific drilling vs. TBI


tunnel and 0 in a rectangular tunnel. The ratio should lie between 0 and 1 in a tunnel

where the roof is arch shaped.

The specific drilling, adjusted after the spacing to burden ratio of the periphery

holes and shape factor, has been related in Figure 5. Consequently, the specific dril-

ling predictive model is derived in Equation 9.

bs ¼ 4:79TBI0:6

m0:5dcþ sh; m=m

3 ð9Þ

5.1. ADJUSTMENTS FOR JOINT ORIENTATION

During field investigations, the joint orientation was found to influence the over-

break/underbreak and face advance which is equal to the product of the pull and

the depth of a round (Ad) divided by 100. To consider the effect of joint plane orien-

tation, the following adjustments are proposed for the major joint sets observed

during field investigations.

The gentle, moderate and steeply dipping joint planes signify the dip angles as

0–30, 30–60 and 60–90 respectively. Similarly, strikes with respect to tunnel

axis are mentioned as parallel, oblique and across to indicate that the joint strike

intersection angle with the tunnel axis as 0–30, 30–60 and 60–90 respectively.

the adjustments due to joint orientations have been suggested separately for pull

and overbreak, which are measured in particular directions.

Further, adjustments for the angle of cut (a) is also provided as a ratio of depth tocut hole length, which is also equal to sin of a ().The relation between pull, adjusted after joint orientation effect and cut angle,

with TBI is displayed in Figure 6. The model for prediction of pull, developed on

the basis of the relation shown in Figure 6, is shown in Equation 10.

Ar ¼½f1:063ðTBIÞ0:55gðsin aÞ2 þ JOAa�

Ad� 100; percent ð10Þ

In case of roof overbreak/underbreak, where the height was increased due to local

fall, specially near the junctions or where the height was abnormally low due to

underbreak caused by misfires, the observations are rejected. The number of values

rejected in this manner are 20 percent of the total population.

Table 5 Joint orientation adjustments

Joint orientation

Dip

Strike with respect to

tunnel axis

Face advance

adjustment

(JOAa), m

Roof overbreak/underbreak

adjustment (JOAr), m

Steep Parallel �0.6 0.6

Steep Across 0.45 0.03

Gentle Across 0.1 0.05

Moderate Across/oblique 0.05 0.2


The relation between the adjusted roof overbreak/underbreak (Ior), taking into

account the adjustments due not only to joint orientation (JOAr) but also to the tun-

nel shape factor (sh) and contour blasting practice defined by the spacing to burden

ratio of the contour holes (mdc), with TBI is shown in Figure 7. Generally the cou-

pling ratio in the contour holes (Rcc) is kept different than that in production holes to

reduce the stress. In such cases the TBI along the tunnel contour (TBIc) becomes

Figure 7 Adjusted roof overbreak or underbreak vs. TBI

Figure 6 Adjusted pull vs. TBI


different that that in the rest of the tunnel section. The roof overbreak/underbreak

predictive model evolved from Figure 7 is given in Equation 11.

Ior ¼ 0:57� 0:52 lnðTBIcÞ � 0:5sh �JOArmdc

; m ð11Þ

6. Application of the developed models

A D-shaped approach tunnel was excavated by heading and benching method

through basaltic formations under aPumped Storage Scheme. The heading area is

29.74m2. The tunnel passed through different varieties of Deccan Traps like com-

pact basalts, amygdolidal basalts and volcanic breccia.

The roof overbreak in the tunnel due to blasting varied from 0.091m to 0.5m in a

length of 385m. The developed models were used to assess whether the geological

features or the drilling error was responsible for such large variations in the over-

break. The results in a the length between 300–370m chainage in the tunnel was

undertaken for this purpose (CMRI Report, 2002).

The geological conditions in the length under consideration were studied carefully.

Three numbers of joint sets were observed in most of the reviewed length of the tun-

nel. An additional joint set was observed in some locations near 370m chainage. The

joints were found tight and undulating with generally rough to very rough surfaces.

Random joints were found inconsistent. Mixed face multiple geological condition

was not observed in the tunnel.

The rock mass properties and the blast results were monitored over a length of

20m strip (covering both right and left sides of the locations) at 300m, 325m and

370m chainage of the tunnel. These three locations represented the tunnel section

between 300 and 370m. The basic properties to estimate TBI in those three locations

are listed in Table 6.

The observed blast results (Obs) in those three locations and the predicted results

(Pred) using the developed models are shown in Figure 8.

It can be seen in Figure 8 that the deviation between the observed and predicted

specific charge and overbreak are not large in between 300 and 326m despite varia-

tions in the formations. However, the predicted overbreak is much in 370m in

comparison to the observed one. This may be due the presence of an additional joint

set in this region.

It appears that not only the joint spacing, which as been taken into account for

RQD estimation, but also the number of joint sets may need to be considered for

more precise prediction of overbreak.

However, it was concluded that the overbreak was caused mostly due to the geo-

logical features and blast design and not because of drilling error. But, the need of

controlled blasting including decoupling between the explosive and blast hole (Rc)

and proper spacing to burden ratio at the periphery holes (mdc) was stressed to

control overbreak.


Table6BasicpropertiesforestimatingTBIintheapproachtunnel

A)Rockproperties(RF)

Locations(m)

RQD

c p(km/s)

nJointorientationwithrespecttotunnelaxis

300

90

5.7

Nil

Dipmoderateandstrikeatobliquetothe

tunnelaxis

326

60

3.1

Dipsteepandstrikeparalleltotunnelaxis

370

88.6

4.4

Dipsteepandstrikeparalleltotunnelaxis

B)Tunnelheadingconfigurations(TF)

Tunnelshape

Heading

area(m2)

Width(m)

Radiusof

curvature(m)

Vertical

wallheight(m)

Tunnel

direction

s h

D-type

29.74

73.5

1.5

Nearly

horizontal

1

C)Blastdesignconfigurations(BF)

Typeofcut

Ca,

Blast

hole

dia.(mm)

Explosive

dia(mm)

Rc

Peripheral

blasting

mdc

Production

holes

Contour

holes

Wedge

60

33

25

0.76

0.76

Contour

blasting

1


6. Conclusions and discussions

As the available models for prediction of blast results do not have a holistic

approach, an index called Tunnel Blasting Index (TBI) has been developed by the

authors to represent tunnel blast environment. TBI includes the most important

parameters influencing the blast results. The blast results like specific charge, specific

drilling, pull ratio and roof overbreak observed during various rounds in four tun-

nels could be well related with TBI. The effects due to tunnel shape and joint orien-

tations, which were not considered in the models developed by others, have been

taken into account in the newly developed relations by the authors.

The predictive models developed in the present paper by the authors are based on

limited site investigations and do not cover many of the possible combinations of the

non-controllable (specially joint orientations), the semi-controllable and the control-

lable parameters. Further, the said models are not dimensionally balanced and are

obtained from near optimised case studies having low to medium advance upto

3m. Hence, the applicability of the models are limited to medium pull only. The

detailed explosive properties and the charge concentration parameters have been

overlooked during field investigations (Section 3) due to lack of variations in the

explosive type and size, and the blast hole diameter in the fields where the investiga-

tions were conducted. The variation in delay timing has not been addressed in the

presently developed models as long delays of more than 100ms were used in most

of the monitored blasting rounds to account for tunnel confinement and wall

damage reduction. Hence, there is a scope to add the variation effects of long delay

time in the models. Though the wall overbreak is not as seriously considered by the

field engineers as the roof overbreak, an effort was made to develop a model for wall

overbreak prediction. However, no acceptable correlation could be established prob-

ably because the overbreak data on both the sides were not collected separately. The

developed models were used in a tunnel to indicate the overbreak possibly occurred

due to geological features and conventional blast design. Further, the control of the

number of joint sets on overbreak was observed.

Figure 8 Observed vs. predicted blast results at various locations in approach tunnel


Acknowledgements

The authors are thankful to Director, Central Mining Research Institute for permis-

sion to publish the paper. Thanks are due to the authorities of the mines and tunnels

where the field investigations were conducted.

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