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Ganda M. Simangunsong

Fakultas Teknik Pertambangan & Perminyakan ITB

6. Underground Blast Design


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Introduction

• There are two reasons to go underground andexcavate:

1. to use the excavated space, e.g. for storage,transport etc.

2. to use the excavated material, e.g. mining andquarrying operations.

• In both cases tunneling forms an integral part of theentire operation.

• The main difference between tunnel blasting andbench blasting is that tunnel blasting is done towardsone free surface while bench blasting is done towardstwo or more free surface.


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Introduction (cont.)

Underground vs. Open Pit Blast Design:

– One free face – explosive consumption is higher thanopen pit blast design

– Environmental constraint – toxic fumes

– Small burden at area of Cut – „Desensitization 75-200 mm; and sympathetic detonation


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Open Pit Blast Design ?


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Blasting Round


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Look Out

Empirical guidance

L = 10 cm + 3 cm/m X Hole depth


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CUT

Various types of Cut (Langefors & Kihlström, 1978)

• Wedge cut or V-cut

• Pyramid or diamond cut

• Drag cut

• Fan cut

• Burn cut


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Wedge Cut or V Cut

• Blasthole are drilled at an angle to theface in a uniform wedge formation sothat the axis of symmetry is at thecentre line of the face.

• The cut displaces a wedge of rock out

of the face in the initial blast and thiswedge is widened to the full width ofthe drift in subsequent blasts, eachblast being fired with detonators ofsuitable delay time.

• This type of cut is particularly suitedto large size drifts, which have welllaminated or fissured rocks. Holeplacement should be carefullypreplanned and the alignment of each

hole should be accurately drilled.


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Pyramid or Diamond Cut

• The pyramid or diamondcut is a variation of thewedge cut where theblastholes for the initial

cavity may have a line ofsymmetry alonghorizontal axis as well as

the vertical axis


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Drag Cut

• The drag cut is particularly suitable in small sectionaldrifts where a pull of up to 1 m is very useful


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Fan Cut

• The fan cut is one-halfof a wedge cut and isapplicable mainly where

only one machine isemployed in a narrowdrive.

• Generally the depth ofpull obtainable is limited

to 1.5 m


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Burn Cut•

A series of parallel holes aredrilled closely spaced at rightangles to the face. One hole ormore at the centre of the faceare uncharged. This is called theburn cut.

• The uncharged holes are often oflarger diameter than the chargedholes and form zones ofweakness that assist theadjacent

charged holes in breaking out theground.

• Since all holes are at right angles to the face, hole placementand alignment are easier than in other types of cuts. The burncut is particularly suitable for use in massive rock such as

granite, basalt etc.


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Rock factor, c

Q = 0.3 kg/m3

C = 1.2Q = 0.36

Example

Ø= 32 mm

B

1.3BB=K= 0.5-1.0 m

0-1 m

The rock constant is an empirical measure of the

amount of explosive needed for loosening 1 m3 of rock


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Rock Factor, c

c?

t?

RMR ?

C = 0.50 + 2.60( t/ c)0.5 + 13 t/ c, kg/m

3

For first trial C=0.4 kg/m3


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Process (cuts)

B1

A1

B2

A2• Define reamer/uncharged hole

• Define 1st cut

• Define 2nd cut andothers


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Process (1st Cut)

6. Define stemming length, hs

hs = 10 d

7. End


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Process (2nd Cut …)

1. Define opening width from the 1st cut, A1

A1 = 20.5 B1

2. Define 2nd burden, B2

B2 = 8.8 x 10-2 [(A1 l WSRANFO)0.5]/ d c

3. Define stemming, hs

hs = 10 d

4. Sub End

5. Repeat point 1 to 3 to find geometry of 3rd cut andso on until the opening width is less than the squareroot of advance (=I0.5) or Burden stoping/lifter .

6. End


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Process (Lifters)

1. Define corrected rock factor, c‟

c‟ = c + 0.05 …… Bn>1.4

c‟ = c + 0.07/B …… Bn


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Process (Stoping)1. Define corrected rock factor, c‟

c‟ = c + 0.05 …… Bn>1.4

c‟ = c + 0.07/B …… Bn


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Process (Contours)

If normal, follow geometry of the downward stoping

If smooth blasting is required, :

1. Define spacing, S

S = (15-16) d

2. Defind burden, B

B = S/0.8

3. Defind charge concentration, l l = 90 d2

4. No stemming is required (fully charged)

5. End


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Hole types Numberof holes

Charge perhole

Totalcharge

1st cut

2nd cut

3rd-4th cut

Lifters

Roof

Wall

Stoping

4

4

8

5

8

6

5

1.59

2.62

3.76

3.20

1.77

3.20

3.20

6.37

10.48

29.36

16.00

14.16

19.20

16.00

Total charge

Opening areaAdvance

Specific charge

Total number of holes

Hole depth

Specific drilling

111.6 kg

19.5 m23.0 m

1.9 kg/m3

40

3.2 m

2.2 m/m3


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Faktor Perencanaan Cut

• Diameter lubang besar (kosong)

• Burden

• Charge concentration

•

Ketepatan pemboran, terutama untuk lubang-lubangledak paling dekat dengan lubang besar/kosong

• Bila menggunakan beberapa lubang kosong, hitungdahulu diameter lubang samaran (fictious diameter )

D = d√n

• D = diameter lubang samaran

• d = diameter lubang kosong

• n = jumlah lubang


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Kemajuan Per Round


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Perhitungan

• Agar peledakan berhasil dengan baik (cleaned blast ),

jarak antara lubang ledak dengan lubang kosong,

tidak boleh lebih besar daripada 1,5 lubang kosong.

• Apabila jaraknya lebih besar hanya akan menimbulkan

kerusakan (breakage) dan jika jaraknya terlalu dekat

ada kemungkinan lubang ledak bertemu denganlubang besar kosong


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a < 1,5 lubang kosong – cleaned blast

a > 1,5 - kerusakan breakage

a 50 ms


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Jarak lubang tembak ke lubangkosong

a = 1,5• a = jarak antara titik pusat lingkaran lubang besar

dengan lubang tembak• b = diameter lubang besar

• Jika gunakan beberapa lubang kosong,

a = 1,5 D• D = diameter samaran


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Pemuatan Lubang Tembak DalamBujursangkar Pertama

• Muatan BP (charge concentration) sedikit → batuan

tidak akan terbongkar.

• Muatan BP banyak tidak akan terjadi blow out melalui

lubang kosong sehingga terjadi pemadatan kembalibatuan yang telah terpecahkan dan efisiensi

kemajuan rendah.

• Kebutuhan muatan BP untuk berbagai jarak C-C

(pusat ke pusat) antara lubang kosong dan lubangtembak terdekat dapat dihitung menggunakan grafik

berikut


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Muatan BP Fungsi Jarak Pusat – PusatLubang Untuk Berbagai Diameter Lubang


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Perhitungan Untuk BujursangkarSelanjutnya

• Perhitungan bujursangkar dalam cut yang tersisa sama denganbujursangkar pertama. Peledakan pada bujursangkar sisamengarah ke bukaan segiempat bukan bukaan sirkular. Sudutledakan (angle of break ) jangan terlalu kecil.

• Dalam perhitungan “burden” (B) sama dengan lebar (W) dari

bukaan:

B = W

• Dengan grafik perkirakan muatan bahan peledak minimum danburden maksimum untuk bermacam-macam lebar bukaan.Muatan bahan peledak ini adalah muatan untuk semua kolom

lubang tembak.• Apabila diperlukan peledakan pada bagian dasar yang susah

diledakkan (constricted bottom) harus digunakan muatan dasaryang besarnya dua kali charge concentration (lc) dan tingginya1,5 B.


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Muatan Fungsi Burden MaksimumUntuk Berbagai Lebar Bukaan


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Stemming Cut

• Panjang kolom lubang bor yang tidak diisi bahan peledak.

ho = 0,5 B

•

Perhitungan berikut utk lubang tembak 38 mm …


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Merencanakan Cut

Bujursangkar I

• a = 1,5

• W1 = a √2

mm 76 89 102 127 159

a mm 110 130 150 190 230

W1 mm 150 180 210 270 320

Bujursangkar II

• B1 = W1

• C – C = 1,5 W1

• W2 = 1,5 W1 √2

mm 76 89 102 127 159

W1 mm 150 180 210 270 320

C-C mm 225 270 310 400 480

W2 mm 320 380 440 560 670


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Merencanakan Cut

Bujursangkar III

• B2 = W2

• C – C = 1,5 W2

• W3 = 1,5 W2 √2

mm 76 89 102 127 159

W2 mm 320 380 440 560 670

C – C 480 570 660 840 1.000

W3 mm 670 800 930 1.180 1.400

Bujursangkar IV

• B3 = W 3

• C – C = 1,5 W3

• W4 = 1,5 W3 √2

mm 76 89 102 127 159

W3 mm 320 380 440 560 670C – C 480 570 660 840 1.000

W4 mm 670 800 930 1.180 1.400


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Geometri Bujursangkar


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Round Stoping & Contour

• Lubang lantai (floor holes)

• Lubang dinding (wall holes)

• Lubang atap (roof holes)

• Lubang stoping arah pemecahan ke atas danhorisontal

• Lubang stoping arah pemecahan ke bawah

• Untuk menghitung burden (B) dan muatan untuk

bermacam-macam bagian dari round dan Contour dapat dipakai grafik berikut …

Burden Fungsi Muatan BP Pada Berbagai


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Burden Fungsi Muatan BP Pada BerbagaiDiameter Lubang Tembak & Jenis BP

G t i P b & P l d k R d


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Geometri Pemboran & Peledakan Round -Normal Profile Blasting

Part of time

round

Burden

(m)

Spacing

(m)

Height bottom

charge

(m)

Charge concentration

Stemming

(m)Bottom

(kg/m)

(Column)

(kg/m)

Floor 1 x B 1.1 x B 1/3 x H lb 1.0 x lb 0.2 x B

Wall 0.9 x B 1.1 x B 1/6 x H lb 0.4 x lb 0.5 x B

Roof 0.9 x B 1.1 x B 1/6 x H lb 0.3 x lb 0.5 x B

Stoping :

Upwards 1 x B 1.1 x B 1/3 x H lb 0.5 x lb 0.5 x B

Horisontal 1 x B 1.1 x B 1/3 x H lb 0.5 x lb 0.5 x B

Downwards 1 x B 1.2 x B 1/3 x H lb 0.5 x lb 0.5 x B


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Number of Holes

Check – Recheck!

Sumber: USACE, 1997


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Specific Charge

Check – Recheck!

Sumber: USACE, 1997


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Initiation Sequence

• The Cut is fired using millisecond delay. Since the rockremoved between each hole in the cut and the centralempty hole must be blown out to provide expansion roomfor the rock removed by the next charge, a long enoughinterval between these holes is needed.

• The recommended delay is 75 - 100 millliseconds.

• 1st and 2nd cuts are recommended to adopt one dalayper hole. Other holes are fired with one delay for twoholes.

• Stoping holes are recommended to adopt 100 – 500milliseconds delay.


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Typical Initiation (cont.)


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Initiation Sequence (cont.)


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Shafts

• Shafts are either driven downwards, sink

shafts/shaft sinking, or upward, raise shafts.


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Shaft Sinking

• Shaft sinking is one of the most difficult and risky blasting jobs as the work area is normally wet, narrow, and noisy.Furthermore, the drilling and blasting crews are exposed tofalling objects.

• The advance is slow as the rock has to be removed

between each blast with special equipment which haslimited capacity. The blasted rock must be well fragmentedto suit the excavation equipment.

• The design of the cross section of the shaft principallydepends on the quality of the rock. Nowadays most of the

shaft are made with circular cross section which givesbetter distribution of the rock pressure, thus decreasing theneed for reinforcement, especially in deep shafts.


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Shaft Sinking

• An important requirement in shaftsinking is to provide optimumfragmentation of the rock so thatit can be cleared quickly from the

congested shaft-face area.• Blasting operation is carried out

against gravity, and the scatter of

the broken rock is confined in theshaft. It is common to use

generous distribution ofexplosives hroughout the rockusing a large number of smalldiameter (35 – 42 mm) shotholes.


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Shaft Sinkingwith benching method

• The most common drilling &blasting methods are benchingand blasting with pyramid cut.

• The benching method is fastand efficient method as the

time-consuming cleaning of thefloor between blast can beminimized. It is easy also tokeep the shaft free from wateras a pump can always be

placed in the lower blasted partof the shaft. The drilling &charging pattern is similar tothat of smaller surface blasting.

Sh f Si ki


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Shaft Sinkingwith pyramid cut

• Shaft sinking with pyramidcut is similar to tunnelblasting with V-cuts. Thedrilling is done with a drillring which is composed of

a circular I-beam to whichthe drilling machines arefixed. The drilling may befixed to the shaft wallswith bolts. Due to

construction of the drillring, the cut will beconical.

Sh f Si ki


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Shaft Sinkingwith pyramid cut

• The number of holesN required for sinkinga shaft of crosssectional area A in m2

is given by:N = 2.5A + 22

• The drilling patternsfor shaft sinking arebasically the same as

those used intunneling butgenerally the cone cutis favoured


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Shaft Sinking

• The explosives used in shaft sinking must always bewater resistant. Even if the ground is dry, the flushingwater from the drilling machines will always stay inthe blastholes.

• The powder factor in shaft sinking is rather high,ranging from 2.0 kg/m3 to 4.0 kg/m3 (Olofsson).

• Nonel type detonators are increasingly preferred forinitiation.


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Raise Shafts

• Older methods:

– Timbered shafts

– Open shafts – up to 25 m, not recommended

• Modern methods:

– Boliden elevator type Jora

– Alimak Raise Climber

– Longhole drilling


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Timbered Shaft

• The rise shaft is driven verticallyand divided into two sections bya timber wall which is extendedbefore each blast.

• When the round is fired, one

section is filled with rock. Theblasted rock will then act as aworking platform for the nextround.

• The second section is used as

ladder-way and fortransportation of equipment, drillsteel, explosive and timber. Theventilation is also placed which iscovered during blasting.


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Timbered Shaft

• Timbered rise shafts havebeen driven up close to 100m, but normally themaximum height should notexceed 60 m.

• The cross section area isusually 4 m2, and theadvance per round approx.2.2 m.


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JORA lift method

• 1950‟s Boliden ABdeveloped the JORA lift.

• A large hole, diameter 110to 150 mm, is drilled froman upper level in the center

of the intended shaft.

• The drilling and chargingare carried out fromplatform on the top of thelift cage and some scaling

can be done from the cagewith protection of theplatform.


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JORA lift method

• The large hole is used as cuthole in the blasting of theround.

• Due to the large size of thecut hole, advances of up to4 m are obtained.

• The area is approx. 4 m2.

• Maximum height is 100 m.

• Free space above the shaft

is needed for the drilling ofthe large hole and forplacing of the lifting gear.


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ALIMAK raise climber

• Introduced in 1957.

• The equipment consists of a riseclimber with a working platform,which cover practically theentire area of the shaft. Under

the platform there is a cage fortransport of personnel, materialand equipment.

• The rise climber is propelled bya rack and pinion system along

a special guide rail.• The rail system incorporated a

tube system for the air andwater supply of the drillingequipment.


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ALIMAK raise climber

• The platform is equipped with aprotective roof under which theblaster stands during scaling anddrilling operation.

• The air driven raise climber maybe used for up to 150 m shaftlength, electric drive up to 900 m.For longer shafts diesel-hydraulicdriven climber are used.

•

The area normally 4 m2

, butinclined (60o) shafts have beendriven full face up to 36 m2.

• The long term advance is approx.3.5 m/day or 70-90 m per month.

Sh ft i i


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Shaft raisingby long hole drilling

• Drilling is done downwards with parallel holes and thewhole are is drilled at the same time.

• Great precision in drilling and charging is a must and thelack of precision has earlier limited the practical height to25 – 30 m.

• Safe - All drilling and charging work is carried out from safelocation.

• Two different cut are used:

– Large hole cut

–

Crater cut

Long hole d illing


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Long hole drillingwith large hole cut

• The large hole cut is still the mostcommon one.

• The drill holes in the round have adiameter of 50 to 75 mm and thecentral large hole is reamed to a

diameter of 102 to 203 mm.• The charging is done from the upper

level. A piece of wood is lowered downon a rope, and when the wood ispasses the lower mouth of the hole

the rope is tightened and the piece ofwood forms a plug for the lower partof the hole.

• The hole should not be stemmed asthe stemming may sinter and block

the hole for the subsequent blast.

Long hole drilling


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Long hole drillingwith crater blasting

• The crater blasting is usedonly for the cut section toopen a hole of approx. 1 m2,then normal stoping will

follow.

• The crater consists of fiveholes, one center hole andfour edge holes. The centerhole is blasted first

whereupon the edge holes areblasted one by one withdifferent delays.

Long hole drilling


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• Before charging the holesare plugged with apiece ofwood, which is lowered downfrom the upper surface on arope and secured to the

lower surface.• The hole is then filled with

sand to the calculated levelof explosive charge. Thecharge should have a

diameter close to that of thehole.

• The charge is then stemmedwith water.

Long hole drilling


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Livingstone‟s theories:

L = 6 x D (mm)

Lopt = 0.5 x Lcrit (mm)

Lcrit = S x Q1/3; (mm);

S : Strain energy (1.5)

Q = (3 x d3 x phi x P)/2; kg;

P : Charging density (kg/liter)

Lopt = 0.5 x S x ((3 x phi x d)/2)1/3 x d x 10 (mm)


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Ring BlastingToe spacing estimation

• Spacing can be estimated as 2 times fractured zone (S = 2 x Rf )

Damage Zone (Hustrulid, 1999)

Ring Blasting


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Blast Damage

• The prediction of damage to the rock mass is a veryimportant factor to evaluate the quality of the excavationprocess in tunneling, so that it would allow theoptimization of explosive charges utilized in successiveblasting rounds, as well as lowering risks of instability

from rock loosening, less support costs and water inflows.• When an explosive charge detonate inside a borehole

several zones can be distinguished in the surroundingrock: 1) Zone of crushing, 2) Zone of radial cracking, 3)Zone of extension and expansion of fractures and 4)

Elastic Zone, where no cracks are formed.


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Excavation Damaged Zone

• Deviations of that perimeter from their outside and insidelimits are called overbreak and underbreak respectively,with the word backbreak used when overbreak isexcessive.

• The factors influencing the magnitude of EDZ can

conveniently be grouped into two categories, which arerock mass characteristics (geological factors) andexplosive (blasting factors)


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Measuring EDZ

There are currently three methods of actually measuringexcavation profiles:

• surveying techniques,

• laser based, and

• photographic light sectioning method (LSM).

The principle of the method is to project a radial light to theperimeter of the underground opening so that light raysintersect the perimeter contour of the cavity. The image of thisperimeter is then saved in digitized form to allow furthercomputerized analysis.


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Measuring EDZ


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Measuring EDZusing PPF and Q

PPF = (Pe.Ee)/Vrp

PPF : Perimeter Powder Factor (PPF)

Pe: Weight of explosive charges used in theperimeter blast holes (kg)

Ee: Unit explosive energy (kcal/kg)

Vrp: Excavated volume of annulus (m3)

M i EDZ


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Measuring EDZ

EDZo : Damage for Overbreak (index)

EDZu : Damage for Underbreak (index)

a, b, c, a´, b´, c´ are coefficients whose values are obtained bymeans of multiple regression statistics correlating Rock MassQuality Q and Perimeter Powder Factor (PPF) with observedOverbreak and Underbreak.

These predictive equations are site specific, but can readily becalibrated to suit other projects, where the rock and blastingconditions differ from the present case.

EDZo = (-a + b.PPF – c.log Q)/100

EDZu = (a´– b´.PPF + c´.log Q)/100

Polynomial 2D

Model First

Order


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Polynomial 2D Model, First order

i2i1oi yb xbb z ˆ

ni

1i

i2

ni

1i

i1o

ni

1i

i yb xbnb z

ni

1i

ii2

ni

1i

2i1

ni

1i

io

ni

1i

ii y xb xb xb z x

ni

1i

2i2

ni

1i

ii1

ni

1i

io

ni

1i

ii yb y xb yb z y

yz

z x

z

b

b

b

y y x y

y x x x

y xn

2

1

o

2

2

Measuring EDZ


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Measuring EDZusing PPV

PPV = k (R/W1/3)-n

PPV : Peak Particle Velocity (mm/s)

R: Distance (m)

W: Weight explosive per delay (kg)

D C it i i PPV


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Damage Criteria using PPV

References PPV (mm/s) Notes

Bauer & Calder (1970) 635 – 2540 Radial cracking/break up rock mass

Langefors & Kuhlstrom (1973) 305 – 610 New cracks/Fall in tunnels

Oriard (1982) 635 Most rock mass damaged

Rustan et. al. (1985) 1000 - 3000 Rock damageMeyer & Dunn (1995) 600 Major damage

Bogdanhoff (1996) 2000 - 2500 Tunnel damaged

Murthy & Dey (2003) 2050 Tunnel damaged (basaltic formation)

Or use tensile strength?t = C PPV

Damage Criteria using PPV in


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g gUG Coal Mines

(a) Major damage: fall of rock/coal blocks from roof and/or pillars.(b) Minor damage: detachment of loosened chips from roof and/or pillars.(c) No damage: no visual damage.


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