Use of Pre-splitting Technique as an Alternative Approach To

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Use of Pre splitting Technique as an Alternative Approach to

Cost Optimisation Tunnel Excavation

T Siimer

1

and H A Giiveni 2

ABSTRACT

Cost of Drilling

32

This paper summarises a cost optirnisation procedure by minimising the

total dril l metres and the overbreak that occurs in tunnel excavat ion in

medium strength rock (uniaxial compressive strength <80Mpa). The cost

optirnisation approach discussed here

is

an alternative interpretation of

the concept of rock fragmentation through seismic waves (Grand, 1980).

The proposed pract ice is an alt ernative approach to smooth blasting

techniques employed in tunnel excavation, since this approach reduces

overbreak and speeds up the excavation. This approach was successfully

employed in derivat ion, energy. and grouting tunnel excavat ions of

Bekrne Dam Project.

The on-site experience and the points that should be laken into account

in pract ice are g iven in the conclusions. The proposed practice was

employed in strong rock conditions (calcareous limestone) at Diyarbakir

irrigation tunnels

in

Turkey with a very limited success.

Cost of Explosives

32

Cost of supports

25

Cost of

transportation

11

INTRODUCTION

The perimeter and face stability of any tunnel excavation

is

dependent on the techniques employed in using explosives as

much

as

it depends on the existing lithology of the rock mass

in

which the excavation is to be carried out. Smooth blasting

techniques and uncontrolled blasting practices not only causes in

stability but also excessive overbreak in soft and medium strength

rock conditions. This overbreak is reflected as additional cost to

contractors that cannot be included

in

the monthly invoices even

though geological conditions are introduced as the main causes.

We have also found out that in environments with fluctuating

humidity conditions the organic

fill

materials that are used to

fill

overbreak volumes and the cavities in the tunnel perimeter cause

the generation

of

poisonous gasses that are flammable. This

increases the cost

of

maintenance and causes insurance problems.

In order to find an economical solution to the existing problem,

pre-splitting was employed first at the tunnel portal then at the

faces. The next step was to optimise the drilling meterage by

eliminating reaming holes that were previously designed to be

drilled as 102 mm in diameter, and two

in

number. The third step

was to reduce the total drill metres on each face by increasing the

drilling diameter from 42 mm up to 89 mm We have

experimented with different set-ups in each step. After six

unsuccessful iterative attempts a unique drill pattern was formed.

The details of this practice and the cost comparisons with

classical drilling and blasting practices are given in the following

sections.

DETERMINATIONOF COST PARAMETERS

In order to form the base

of

this practice a research was

performed in 12 different tunnels with face areas of 20-75 m

2

and where drilling and blasting practice

is

employed with

drillholes ranging from

8 45

mm in diameter. The above

practice

as

here in after will be named

as

lassical approach . All

the tunnels that are surveyed was chosen to be in calcareous rock

formation in the presence

of

at least one fault line.

I. Department

of

Tendering and Planning, Ceylan Construction

Co Ltd, Ankara, Thrkey.

2 Department of Computer Engineering and Information Science,

Bilkent University, 06533 Ankara, Thrkey.

FIG I - Cost distribution for tunnel excavation.

The result

of

excavation cost (manpower included) calculations

of these tunnels that are excavated or are being excavated can be

summarised

as

in Figure I (Stimer and Asenjo, 1993).

In order to give a comparison

in

open cast operations where

bore hole diameters

of

64-89 mm is used (in calcareous

environment), the cost of drilling only is approximately

1.30 1.60 /m and for each drill metre approximately 12 m

3

of

rock can be excavated. On the contrary in tunnel excavation the

drilling cost is approximately 2.60-3.20 m and only

0 6Q O 90

m

3

of

rock can be excavated per drill metres.

In classical approach, after a few experimental blasts the

required amount

of

explosives needed for each face blast can

easy be found depending on the existing rock media of concern.

Then modifications to theoretical calculations can easily be made

to suit the local boundary conditions. Therefore, if one does not

take into account the drastic changes

in

the lithology of the rock

environment, the amount of explosives required for each face

blast can be assumed to be approximately fixed. Trying to cut

down the amount

of

explosives required to a further degree will

decrease the excavation speed and increase the maintenance cost

of

equipments used.

The number of workers needed for excavation depends totally

on the past experience of the engineer

in

charge. Therefore the

required work force will be fixed in number, depending on the

rate

of

excavation planned and number

of

shifts practiced. Trying

to reduce the number

of

worker will require high mechanisation

that will bring an additional inve tment, hence the decision on the

degree of mechanisation will be made on the basis of local

conditions.

The cost of supports depends only on the geological conditions

that will be faced during excavation, since research drilling

cannot be made on every metre of the tunnel alignment.

Therefore the cost of supports depends on the amount

of

risk that

will be planned to be undertaken by the contractor. The risk

is

very limited due to the human life involved. As a result, cost

of

support can also be considered to be fixed.

We therefore are left with only one cost parameter that

is

the

drilling cost. This parameter

is

interrelated with explosive cost.

We

all know that

no

one can blast effectively a wrong drilled

tunnel

face

We

have encountered tunnel face pulls less then 80 per

cent of the drilling bore hole length due to miss practices

in

drilling.

EXPLO 95 Conference

Brisbane, 4 - 7 September 1995

383



T SUMER AND H A GUVENIR

PRELIMINARY COST OPTIMISATION

Figure 1 shows the distribution of costs for tunnel excavation.

In

order to optimise the drilling cost,

we

programmed on-site control

of the drilling activities, considering the tunnel excavation studies

we

have studied previously. This control saved us

8-12

per cent

on overall drilling costs. As the second step

of

cost optimisation,

various drill patterns for pre-splitting in tunnels are trailed. Six

out

of

23 tunnel face blasts were completely a failure. We then

were able to achieve partial success in the following

13

blasts. In

the last four blasts

we

achieved full success.

PRE-SPLITTING PRACTICE

The

theoretical background which was modified due to present

boundary conditions of this pre-splitting practice was based on

the assumpt ion that in order to form a crack pattern within the

rock mass, the vibrations thal will be generated from the blast

of

the explosives tha t are placed on the per imeter should

be

kept

within the limits given in Table

1.

T LE

1

Critical vibration velocity

r

rock conditions

ROCK TYPE CRITICAL VIBRATION

VELOCITY mrnIsec

A Hard Rock

Density >2.8 ton/m

3

>950

u >220MPa

B

Medium Hard Rock

Density >2.6 ton/m

3

600--750

70 <

u

< 140MPa

C Soft Rock

Density> 2.3 ton/m

3

<400

u

<40MPa

Note:

u

Uniaxial eompressive Strength

The calculation of the expected vibrations can be done

Johansson, 1990 as follows:

V

=

YIE *C

where V

=

Critical vibration velocity mm/ sec,

Y

=

Rock Strength DCS ,

E = Young Modules,

C = P wave velocity mm/sec.

A study performed by Berta 1990 also gives a practical

background.

PRE-SPLITTING

AT THE

PORTAL

An example of pre-splitting practice employed at the portal of a

tunnel is shown in Figure 2 Popouic and Sumer, 1989 .

Pre-splitting holes of diameter 34--45

mm

are drilled at the

boundary of the portal with 20-35 cm spacing and loaded with

counter explosives as shown in Table 2 .

The length

of

these

pre-splitting holes are extended as the RQD Rock Quality

Designation of the rock gets higher, this ranges from 1.5 m to

5 m in depth. A simi lar approach w as mentioned Mavar 1987

two years after this application was realised. These mentioned

counter holes which were drilled and blasted before drilling of the

production holes.

PRE-SPLITTING AT THE TUNNELFACE

The idea of optimisation of drilled metres in tunnels occurred

when we placed a few 60 mm gelatin cartr idges in

one

of

the 102

mm reaming holes after experiencing several problem with

scaling and blasting.

No. of holes: 117

empty:

loaded: S9

FIG

2 - Pre-splitting technique applied at the portal of energy tunnel

number 2.

T LE 2

Hole spacing hole diameter and explosives used

r

various rock

conditions at tunnel portals

Type of Rock Hole S pacing Hole D iameter Explosives

UCS MPa cm mm

Used

Shale-soft-limest 20 cm 45-49 GURITBI7

ne

Kountur Vitezit

UeS:30-40 5a 22

Dolamitic 22

43 GURITAII,17,

Limestone Kinit20

ues:

60-80

Massive 23-27 41

GURIT BII,17,

Limestone

Kontur Vitezit

ues: 80-110

18

Note:

u

Uniaxial eompressive Strength.

We found out that scaling is no more a problem. This inspired

the idea

of

increasing the blast hole diameters from

38-45

mm to

52- j4-76 mm so that we can load fewer holes with a larger

amount of explosives per drilled hole. In the first trial only the

drillhole diameters are increased and a reasonable reduction in

total dril ling meters was achieved, but excessive overbreaks

occurred. In order to reduce the extent of overbreak, the perimeter

of the tunnel face was drilled with pre-splitting holes 50-70 cm

apart. These pre-spitting holes were drilled parallel to the axis of

the tunnel and loaded with explosives of 4300 m/sec detonating

velocity. The loading density of these holes was about

0.2-0.35 kg/m. These perimeter holes were blasted with 25 m/sec

delay per 15-25 holes Fourhaugh and Sumer, 1989 , in order to

bring down the blast vibrations generated from pre-splitting to the

calculated limit. This enabled us to form a crack all a round the

volume

of

rock tha t is planned to

be

blasted in the next round.

The

next step was to drill the face that was previously separated

from the rock medium with 52-64 mm holes.

This saved us 25 per cent in dril ling cost. Another ten per cent

saving was obtained from blasting since higher loading densities

were achieved by employing large diameter holes. The last saving

was from the over break which was 65 per cent.

Another prob lem that we have encountered was the uneven

tunnel face generated after each blast. This problem was

overcome by increasing the loading density

of

explosives

gradually towards the end of the blast holes. By this practice we

have found out that we can widen the tip angle of the breaking

cone that is formed at the far end of the blast hole. We have seen

that the breaking angle totally depends on the velocity of

detonation and density of explosives loaded at the tip point. After

realising this fact, a piece

of

gelatin explosive is placed at the end

of each pre-splitting holes Figure 3 .

384

Brisbane, 4 - 7 September 1995

EXPLO 95 Conference



~ J ~ ~ ~ f ~ ' ~ U § ' ~ ' : ' : ' ~ r E ' ; ~ '

' ~ 4 \ i E ' ' ' ~ j ~ § ~ § , , ~ w ~ 1 ~ ~ / 2

lllJLIT

ISO

1 2 EIlULIT ISO Sxl GURU 32/1100

,

40 200

, 17/ 500

FIG

3 - Loading detail

of

a pre-split hole.

FIG 4 - Classical burn hole drilling detail.

USE

OF

PRE-SPLITIING TECHNIQUE

In order to increase the savings from drilling costs to a further

level, the design of the classical reaming hole (Figure 4 is

changed. This idea

is

inspired by Bergovists (1990). Seven to ten

holes were drilled parallel to the tunnel axis with diameter of

43 mm (Figure 5). These holes were loaded with contour

explosives and were connected with detonating cord P5 in a

vertical direction and blasted with the delay pattern shown in

Figure 5. Three parallel crack lines were formed by this method

at just the same location

as

the previous bum cut used

to

be

drilled. Through this application, in place

of

two 102 mm holes

only ten 43 mm holes were drilled which took only 25 minutes

instead of 1.5 hours, partly due to the need to change drill bits

and related accessories. This itselfcaused a saving of 66 per cent.

The next step for cost optimisation was to increase the drillhole

lengths

as

we enter a rock media of RQD 85. The drillhole length

of 3.2 m was increased to 5 m so a pull

of

m was

realised. Through this application the number of faces that needed

to be blasted was decreased by up to 40 per cent. Even longer

drillholes were experimented with by use of couplings. Due to the

lack of mechanisation (which is now available in the market

at

a

very reasonable price) and due to the complexity involved in the

explosive loading and blasting patterns of longer blast holes, the

cost

of

tolerable mistakes was so high that we could not claim to

have reached a successful result.

We

believe that longer blast hole

drilling in medium strength rock conditions needs further

research.

The results of the application of all the techniques mentioned

above at the same tunnel face resulted with the following savings:

I. Reduction of overbreak was 60 72 per cent. Reduction in

support cost was 25-35 per cent. Reduction

in

shotcrete was

40 50 per cent.

2.

Reduction

of

the fill concrete cost was around 72-81 per

cent.

3.

The time gain helped

us

to complete an excavation round

within a shift cycle.

rts

trial

Second trial

THE COST N LYSIS OF THE PRE SPLITTING

PPRO CH

~

~ No

Last

trial

No

FIG 5 - Alternative

bum

hole drilling pattern.

For a tunnel face with a face area of 40 m

2

and a circumference

of 25 m a pull of 2.6 m per blast was practiced. Since the classical

approach was used, 71 production holes were drilled per face.

The cost of drilling was 71 holes 3.2 m per hole 3

Im

681.6 per face. When the pre-splitting approach was employed

45 holes were enough to pull the face as the first attempt so the

drilling cost was 45 holes 3.2 m 3.4 Im 489.6 per face.

This means a saving of 191.9 per face. The classical approach

causes

an

overbreak minimum 12 cm, which results in extra

0.1

m 25 m

=

3 m of concrete per metre. The cost incurred

is

3 m Im

65

3

195 Im

In

the pre-splitting approach 60 per

cent of this loss was regained resulting 195 Im 0.60 117 Im

of

saving per face. In the classical approach I ton/m contact

grouting was required however in the pre-splitting approach,

0.70-0.83 ton/m grout was enough. The saving in grouting was

21

to (1-0.83)

=

35.7 per metre of tunnel.

As

a result, a

savings of 343.37 per metre of tunnel wa achieved. Please note

that the savings discussed above are in terms of only three cost

parameters. When all the other cost parameters are considered,

approximately a saving of 510 per metre was calculated. If one

considers that the total cost of tunnel operation is approximately

2500 per metre, the total savings will be up to 20.4 per cent on

cost bases.

PR CTIC L CONSIDER TIONS

Some

of

the important points that are needed

to

be taken care of

in the application

of

pre-splitting approach in weak

to

medium

rock conditions can be listed

as

follows:

EXPLO 5 Conference

Brisbane. 4 - 7 September 1995

385



T SUMER

AND

H A GUVENIR

FIG

6 - The relation between type of linejoining

two

adjacent pre-splitting

holes and the amount of explosives.

straight lines

correct amount

of explosive

* convex lines

* insufficient amount

of

explosive

concave lines

excessive amount

of

explosive

Grand, C

H

1980. An Emprical Method Enegy Distribution in Blast

Patterns Soc of Mining Engineers of AIME).

Johanson, M

1990.

Tunnel Blasting Technique Course in AdvancedRock

Blasting Techniques Section 4 Gytrop: Seweden).

Mavar,

R

1987. Results of Channel and Tunnel Po:lJ . Excavation

Practice, Technical Report

No:

2220-1-223,915-87.

Popouic, R and Sumer, T

1989.

Project Procedure Excavation and

Diversion Tunnel Inlets. Enka Cons Report No: DTOO28/01 LOT

4A

BEKHMEDAM.

Sumer, T and Asenjo, T R 1993. An Interactive Computer Simulation r

Tunnel Cost Calculations Interm Report no: CICS 0028193 Berke

Dam

Project, Ankara.

Sumer, T and Guvenir, H A

1993.

Cost optimization

for

controlled

blasting, transportation, and crushing,

in

Proceedings Nineteenth

Annual Conference on Explosives

and

Blasting Technique pp

13-23

International Societyof Explosives Engineers).

REFERENCES

The diameter of per im eter holes

should

be gradually

inc re ased f rom 38 mm to

mm. The hole spacing should

be inc re ased f rom 35 cm to 70 cm gradually. We have faced

several problems when the spacing of pre-splitting holes

were increased beyond 60

cm

in medium hard rock

conditions.

2. Production

holes

of

diameter up

to 89 mm

should

be placed

at the

middle

of

the

tunne l f ac e

and the hole

diameters

should gradually be decreased near the per im eter to a void

excessive v ib rat io ns an d ov er brea ks . The concept of

changing drillhole diameters in the same drill patte rn was

also experienced in

open

cuts Stimer and Gtivenir, 1993).

3. In

determining

the

spacing

and

loading

density of perimeter

hol es, first a co ns er vat iv e

spacing should be

fixed

and

perimeter explosive should be changed. When the required

amount of explosive per hole is almost determined then the

spacing should be optimised.

4. Utmost care and control on the drill ing pattern and its

compatibility with the calculated drilling patterns should be

maintained on-site.

5. In

order to

reduce

the

cost

of counter

explosives which are

around 4ooo/ton low density

ANFO 0.3-0.5

kg/dm

3

) was

used. The point to be taken care in t he use of

ANFO

is tha t

the absorption of fuel oil sh ou ld n ot b e l es s chan 8. 5 per

cent, otherwise unobserved fuel-oil dissolves the synthetic

spheres

introduced

to the mixture.

6. In medium-hard

rock

conditions detonating cord of P50-60

can

be used instead of specific counter explosives. The

point here is t ha t two lines of

P30

will no t replace

on e

line

ofP60.

7. In any case, all t he counter explosives and explosives in

production holes should be primed from the bottom of the

holes.

8. All the sh ock tubes a nd detonator s should be carefully

checked

by

an

inspector before being sent

to a site.

9. The success of pre-splitting blasting can be e va luated by

examining t he l in e joining

two

adjacent pre-splitting holes

as follows:

if amount of explosive is

normal

then the line is

straight,

if

excessive

amount of explosive is u se d, t he n t he l in e is

concave,

if insufficient amount of explosive is use d, the n the line

is convex.

These cases are illustrated in Figure 6.

Ariog1u E and Tokgoz, N 1993. Kaya\;larin Tek Eksenli Basin\;

Dayanimi Arasinda Ampirik Baglantilar Ozerine Bir Istatikse1

c;aIisma, Geosandjoumal 23.

Bergovists, I

1990.

Tunnel Blasting Lecture Notes. Blasting Application

Research Nitro Nobel ab: Gyttrop Seweden).

Becta

G

1990.

Explosives an Engineering Tool Itallsplosives: Milano)

pp

100-110.

Fourhaug, M and Summer,

T

1990. An Alternative Aproch to Audio Over

Break Problems Experianced at Urfa Tunnels. Sevedofo Consult,

Report no: DS0028116

86

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Use of Pre-splitting Technique as an Alternative Approach To