Determination of Stope Geometry in Jointed Rock Mass

© JCRM All rights reserved.

Volume 5, Number 2, April 2009, pp.63-68

[TECHNICAL NOTES]

Determination of stope geometry in jointed rock mass

at Pongkor Underground Gold Mine

Budi SULISTIANTO*, M. Safrudin SULAIMAN

**, Ridho Kresna WATTIMENA

*,

Achmad ARDIANTO***

& Kikuo MATSUI****

* Member of ISRM: Dept. of Mining Engineering, ITB, Bandung 40132, Indonesia

** Dept. of Mining Engineering, ITB, Bandung 40132, Indonesia

***PT Antam Tbk., Jakarta12530, Indonesia

**** Member of ISRM: Dept. of Earth Resources Eng., Kyushu University, Fukuoka 819-0395, Japan

Received 20 01 2009; accepted 16 04 2009

ABSTRACT

One of the stopes which is being mined by Pongkor Underground Gold Mine using cut-and-fill mining method is located in South

Ciurug Mine. The rock mass in this location is weak due to the extensive joint occurrence. In order to understand the effect of rock

joints to the stability of stope, an investigation was carried out by means of specific drilling from two crosscuts to acquire rock samples

from this location, and observing the joints orientation in the drilled holes wall using a borehole camera.

The stability of stope due to the extensive joints occurrence in this location was then analyzed by empirical and numerical methods.

Rock mass classification in this location was evaluated by RMR and Q-System method. Numerical analysis was carried out by

simulating models which represent the rock mass conditions, using 3DEC Program. The analysis was focused on the back of the stope,

due to the effects of joints existence in this location.

Keywords: Cut-and-fill underground mine, Rockfall, Stope-back, Jointed rock, Borehole camera, Underground stability analysis,

Numerical analysis

1. INTRODUCTION

Instability in an underground mine manifests itself

especially as rockfall from the roof or back of the opening. In an

underground mine located in jointed rock mass, rockfall is

formed in blocks due to the intersection of the joints (Hoek and

Brown, 1980). Rockfalls from unstable stope-back vary in size

from small blocks occurred between support units to huge

blocks as wide as the opening. Pongkor Gold Mining Business Unit which belongs to PT

Antam Tbk is one of the underground gold mining sites which

implements cut-and-fill method. In this method, the ore body

being mined is located in stope, and the access to the stope is

through a crosscut from a ramp. The structure of ore body being

mined is commonly in a form of a vein body, which the

thickness varies from 7m to 15m. The general orientation

(strike/dip) of the vein ore body is N3300E/60-700. The rock

mass condition in which the ore being mined is ranging from

fair to poor rock.

During the mining operation, stope failures (mostly rockfall

from stope-back) happened at certain locations. In one of the

stope located in South Ciurug mine, the rockfall occurred with

size of about 2mx2mx1.5m. The failure seemed to be controlled

by the intensive existence of joint along the ore body.

A special investigation was then carried out to grasp the

influence of the joint existence to the stability of the stope. The

investigation was located in Xcut-10 and Xcut-12 in South

Ciurug Mine. The investigation was carried out in a rock mass

located in a stope which was about to be mined through

boreholes drilled from a Xcut. The result of this investigation

was the position and orientation of each joint in the rock mass

penetrated by investigation boreholes.

The aim of this research is to analyze the stability of stope-

back due to the effect of joints existence based on the

investigation result. The analysis was carried out by means of

empirical and numerical methods. Rock mass classification in

this location was carried out by RMR and Q-System method

(Bieniawski, 1989). Rock support system required to stabilize

the stope-back was than determined based on the rock mass

classification. Numerical analysis was carried out by simulating

models which represent the rock mass condition and support

system, using 3DEC Program.

64 B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp.63-68

2. SITE INVESTIGATION

Location of this research was South Ciurug Mine that was

located in Level 500m. The rock mass in this location was

constituted by two main types of rocks, which are andesite in

sidewall rock (hanging-wall and footwall) and quartz as vein

ore body. The sidewall rock is generally strong and only few

joints were developed inside, while the vein ore body is

classified as poor rock and heavily jointed. Ground water

condition in this research area was dry to dump. Map showing

the research location and typical cross section of lithology in

this location are depicted together in Figure 1.

2.1. Horizontal and Inclined Drilling

Drilling was carried out to grasp the rock mass condition in

front of the crosscut where stope is to be mined. Bore holes

were drilled from two crosscuts within this research location,

which are Xcut-10 and Xcut-12. Two bore holes were drilled in

each crosscut with different orientation as can be seen in Figure

2 and Table 1.

Figure 1. Research location and typical rock lithology

Figure 2. Drilling at research location

2.2. Laboratory Testing

Rock sample specimens obtained from the drilling were

tested in laboratory. The tests included physical and mechanical

properties test of the rock specimens. The mechanical properties

test consisted of Uniaxial Compressive Strength test for intact

rock sample and Direct Shear test for discontinuity plane. All

tests were conducted in Laboratory of Geomechanics,

Department of Mining Engineering ITB. The average values of

each test parameter was then determined, as listed in Table 2.

Table 1. Drilling orientation

Table 2. Physical and mechanical properties of rock

No Material

(gr/cm3)

E

(GPa)

C

(MPa)

(..0)

c

(MPa)

1 Sidewall

Andesite 2.49 18.93 0.23 - - 104.06

2 Quartz

Vein 2.57 11.28 0.26 - - 32.59

3 Joint - - - 0.162 21.5 -

Note : : Rock Density C : Cohession

E : Young’s Modulus : Angle of internal friction

: Poisson’s Ratio σc : UCS

2.3. Rock Quality Designation

Rock Quality Designation (RQD) which describes the rock

mass condition was obtained from core drilled from this

location. Core obtained from the drilling was usually in

fractured condition due to the extensive joint occurrence. The

rock masses in this research were then classified as poor rocks.

This assumption was verified by the RQD values which vary

from 10% to 96% and 15% to 73% for Xcut-10 and Xcut-12

respectively. RQD values obtained from this drilling showed

Figure 3. Drill-core from Bore-hole GTC-02 at X cut-10

DH Location

Orientation

Remarks Depth (m) Trend

(N .. E)

Plunge

( ..°)

GTC 1 Xcut-10 85 -5 Horizontal 15.70

GTC 2 Xcut-10 85 -40 Inclined 17.30

GTC 3 Xcut-12 85 -15 Horizontal 14.50

GTC 4 Xcut-12 85 -42 Inclined 14.05

Joint A

Joint B

Joint C C

B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp.63-68 65

that the rock mass is in range of very poor to excellent condition.

Figure 3 shows the example core condition obtained from this

location.

2.4. Discontinuities Observation

To observe the joint existence along the borehole, special

equipment was used to grasp the joint position and the

orientation. A set of borehole camera apparatus was inserted

into borehole to capture image of joint along borehole. The

borehole camera equipment which was used during the

observation is shown in Figure 4.

Figure 4. Borehole camera apparatus

Each discontinuity images obtained from borehole camera

observation, was treated using the procedure proposed to

determine the orientation of each discontinuity (Sulistianto et.al.,

2007). As the result, the orientations of joint A, joint B, joint C

for example, from Fig. 3 are N1100E/470, N0300E/560 and

N0750E/320 respectively. After all images were processed, and

the orientation of each discontinuity has been calculated, the

results are then plotted in the stereonet for further analysis.

The result of this observation showed that joint condition in

this location consists of three major discontinuity sets plus

random discontinuities both in Xcut-10 and Xcut-12. Figure 5

shows the result of the observation plotted in stereonet. Since

Xcut-12 located relatively in the upper part of Xcut-10 (figure

2), the joints showed in good persistence.

Figure 5. Discontinuity orientation at Xcut-10 (left) and Xcut-

12 (right)

3. STABILITY ANALYSIS OF STOPE-BACK

3.1. Empirical method

Stability analysis using empirical method was conducted to

estimate the support system required for the stope-back. Input

parameters for this calculation were obtained from both site

investigation and rock sample testing.

Based on the investigation carried out in the Xcut-10 and

the Xcut-12, it was revealed that the RMR values of the

andesite rock were 56 and 51 respectively. So, the rock masses

located in the footwall (Figure 1) were classified as fair rock

and will give no problem to the stability (Sulaiman, 2007).

The stability of the stope-back, which will be located in the

vein ore body in front of the Xcuts was approached by the rock

mass classification methods of RMR and Q Systems. Rock

classification was carried out based on the drill holes core data

to predict the rock mass condition in stope-back.

The RMR value of the vein ore body in front of Xcut-10

and Xcut-12 were 35.75 and 30.4, respectively. Therefore, the

vein ore body was classified as poor rock. Meanwhile, based on

Q-system, the obtained Q values were 0.43 and 0.40 for Xcut-

10 and Xcut-12 respectively (Sulaiman, 2007). For ensuring the

stability of the opening, determination of support system was

conducted by using both rock mass classification methods.

From RMR system, it is recommended that systematic bolt 4-

5m long, spaced 1-1.5m and shotcrete with thickness of about

10-15cm should be applied. While Q system suggested that

systematic untensioned bolting spaced 1m and mesh-reinforced

shotcrete with thickness of about 5cm should be applied.

Regarding to the mine opening, however, the recommended

support system by both rock mass classification methods was

considered to be over-confidence. The recommended support is

suitable for the long term usage of underground opening, such

as underground civil construction. In an underground cut and

fill mining, which applies overhand stoping method, stope

which will be mined only act as temporary opening, and only

need simple support system to ensure the opening stability.

After the extraction, the empty stope will be filled with slurry

and then mining operation move to the higher level. More

support installed on the stope-back, would create more

disturbance during mining operation sequence in the upper level.

Consequently, support system which will be used must be as

simple as possible but strong enough to stabilize the opening.

In this underground mining operation, the only support

system which can be accepted in the stope was combination of

friction bolt, wired mesh and W-strap. Bolt spacing obtained

from empirical recommendation and used as rule of thumb in

this area was 1m.

3.2. Numerical method

Numerical method was carried to evaluate the performance

of support system recommended by the empirical method.

Numerical software used in this research was 3DEC

(3Dimensional Distinct Element Code). This software is a three

dimensional software which allow to model the discontinuities

in the rock mass.

The discontinuities in the ore body model are joints

obtained through borehole camera investigation inside two bore

holes from each crosscut. Figure 6 shows the discontinuities

around the vein body located above the stope which were about

to be mined. While the rock mass parameters used in the model

is shown in Table 3.


Support system implemented in the model consisted of axial

bolt and w-strap whereas wire mesh was not included in the

model, due to the limitation of the program. The axial bolts of

2m length were placed vertically at stope-back with spacing of

1m. W-straps were modeled as thin plate which tied the bolts in

each row. Mechanical characteristic of both support are showed

in Table 4 and Table 5. Series of calculation were carried out by

simulating the bolt spacing from 2m to 0.5m. In order to

evaluate the stability of the stope opening, vertical deformation

from particular points (points 1 to 9) located in the stope-back

of the stope were then observed.

Table 3. Rock Parameters used for Numerical Calculation

No Material

(gr/cm3)

E

(GPa)

K

(GPa)

G

(GPa)

1 Sidewall

Andesite 2.49 18.93 0.23 11.69 7.70

2 Quartz

Vein 2.57 11.28 0.26 7.83 4.48

C

(MPa)

(..

0)

Jkn

(GPa/m)

Jks

(GPa/m)

3 Joint 0.162 21.5 1 1 -

Note : : Rock Density C : Cohession

E : Young’s Modulus : Angle of internal friction

: Poisson’s Ratio Jkn : Normal Joint Stiffness

G : Shear Modulus Jks : Shear Joint Stiffness

K : Bulk Modulus

Figure 6. Discontinuity model at x-cut-10 (left) and x-cut-12

(right)

The result showed that vertical movement of the stope-back,

which indicates instability, still occurred on spacing more than

1m. Therefore the bolt spacing was narrowed. Figure 7 shows

the result of numerical modeling using 1m bolt spacing.

By using 1m bolt spacing, the vertical movement was

around 5 cm at stope-back in front of Xcut-10. Even though few

small blocks still moved down, however, catastrophic

movement can be avoided. Initial deformation occurred at the

beginning, as the result of rock-support interaction, however

after reaching computational step of 6000, vertical deformation

tended to be constant, around 5cm at point 7 (see Figure 7 lower

left).

Table 4. Mechanical properties of rock bolt

47 mm Bolt Minimum Typical Yield Strength 345MPa 120kN 445MPa 160kN

Ultimate Tensile Strength of Tube 460MPa 165kN 510MPa 180kN

Friction Bolt Diameter 47mm

Hole Diameter Range 43mm min/45.5mm max

Mass per Meter 2.79kg

Cross Section Area 355mm2

Meanwhile, at a stope-back of stope in front of Xcut-12,

with spacing of 1m, there were blocks which still moved down.

It also can be seen in the vertical deformation graph that some

blocks moved down whereas others tend to reach stability after

giving large movement (see Figure 7 lower right). Though the

number of the loosen blocks are small, but the size of individual

blocks were significant. So it is important to note that the

stability of the opening in this location can be obtained by

reducing the spacing. Another model was developed by 0.75cm

bolt spacing. The result showed that smaller deformation was

obtained (around 5mm at point 3 and 4) as shown in Figure 8.

Table 5. Mechanical properties of w-strap

NOMINAL

WIDTH

TYPICAL WEIGHT

PER METER

MATERIAL

GRADE

TYPICAL YIELD

STRENGTH

mm Kg kN

240 3.78 AS/NZS 1594

Gr.HA350 164

3.3. Determination of optimum stope geometry

Geometry of unsupported stope-back can be estimated from

stability chart developed by Potvin (Huchkinson and Diederichs,

1996) based upon the calculated stability number N’(see

equation 1) and the calculated shape factor or hydraulic radius

(see equation 3).

CBAQN '' ................................................................(1)

a

r

n J

J

J

RQDQ' ................................................................(2)

Where Jn, Jr and Ja were joint set number, joint roughness

number and joint alteration number from Q-system

respectively. While A, B and C were correction factors

which considers the relation of intact rock strength and insitu

stress, the relative angle between dominant discontinuity

orientation and excavation, and the gravity effect to the

stability of excavation surface respectively.

Hydraulic radius of surface analysed = area /perimeter ….. (3)

Using rock parameter used for calculating Q values

mentioned above and considering that excavation is carried out

at around 200m below the surface and also the maximum insitu

stress resulted from measurement is around 4 MPa (Sulistianto

et.al., 2003), the stability number N’ calculation is given in

Table 6.

Table 6. Stability Number (N’) Calculation

Parameter Xcut-10 Xcut-12

Description Rating Description Rating

RQD 51.6 51.6 48 48

Jn Three Sets +

Random 12

Three Sets +

Random 12

Jr Rough, Undulating 3 Rough, Undulating 3

Ja Zone of

disintegrated 12

Zone of

disintegrated 12

Jw Minor inflow 1 Minor inflow 1

SRF Low Stress, Near

Surface 2.5

Low Stress, Near

Surface 2.5

Q RQD/JnxJr/JaxJw/SRF 0.43 RQD/JnxJr/JaxJw/SRF 0.4

Q’ 1.075 1.000

A σc/σ1 = 32.59/4 0.8 σc/σ1 = 32.59/4 0.8

B α = 45° 0.5 α = 45° 0.5

C Stope – back 2 Stope – back 2

N’ 0.86 0.8

B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp.63-68 67

Figure 7. Numerical model result of stope in x-cut-10 and x-cut-12 by using 1m bolt spacing

Figure 8. Vertical deformation at stope-back of x-cut-12 by

using 0.75m bolt spacing

By assuming that the width of the opening was constant at

10m, which was general vein width in this location, and the

height of opening was 4-5m considering the height of Jumbo-

drill, the hydraulic radius of stope-back was then calculated

based upon the simulated length of unsupported excavation

from 4m to 100m and the result were plotted in Potvin stability

chart as can be seen in Figure 9 and Table 7. The result showed

that the stope-back will be in safe condition within area of

10mx4m.

4. DISCUSSION

Based on the result of stability analysis and considering

mining operation which applied overhand cut and fill stoping

method, the support system which is combination of friction


bolt 2m long spaced 1m, W-strap and wire-mesh is still able to

be applied in the level of Xcut-10. However, bolt spacing

should be changed to 0.75m if mining operation in the level of

Xcut-12 is started.

Figure 9. Stability chart of research area

Table 7. Stope condition related to opening geometry

Sto

pe

Wid

th (

m)

Sp

an

Wid

tho

ut

Su

pp

ort

(m)

Hy

dra

uli

c

Ra

diu

s (m

)

Remarks

Xcut 10 Xcut 12

10 4 1.4 Stable Stable

10 8 2.2 Transition Zone Transition Zone







Figure 10. Stope-back is supported by friction bolt, w-strap,

wire-mesh and cribbing

It is obtained from optimum stope geometry calculation

that unsupported stope-back which is still safe to be mined

would be 10mx4m. It is in a good agreement with practical

condition which has excavation progress around 3m span due to

2.4m length of blastholes drilled by Jumbo-drill, and the

opening is in stable condition. Even though stable condition,

after 1 hour smoke-clearing activity the miners do scaling

immediately at the stope-back and then continued by

installation of 2m long of friction bolt using 1m spacing. In

order to protect the rockfall, w-strap and wire mesh is also

installed. For anticipating the inconsistency of vein width due to

mineralization process, timbering system (called Cribbing) is

sometimes applied if the width of vein becomes wider (> 10m)

as shown in Figure 10.

5. CONCLUSION

Rock mass condition in this location, especially in the stope

which was about to be mined was classified as poor rock, due to

extensively developed joint in the rock mass. Suitable support

system which has been confirmed by numerical method and can

be practically used considering the operation constrain would be

2m long frictional bolt with spacing of 1m at Xcut-10 and

0.75m at Xcut-12, w-strap, and wire mesh.

The optimum stope geometry, which is still safe to be

mined, considering the vein rock mass conditions would be of

10m by 4m area of unsupported stope-back during operation.

6. ACKNOWLEDGEMENT

The authors would like to acknowledge PT. Antam, Tbk,

UBPE Pongkor management for the possibility and opportunity

to conduct this study, and Puslitbang TEKMIRA management

for their help using the 3DEC calculation. The authors also

thanks to the members of Geomechanics Laboratory,

Department of Mining Engineering ITB for their help.

REFERENCES

Bieniawski, Z.T., 1989, ”Engineering Rock Mass Classifications”, John

Wiley & Sons, New York.

Hoek, E., and Brown, E.T., 1980, “Underground Excavations in Rock”,

The Institution of Mining and Metallurgy, London.

Huchkinson, D.J. and Diederichs, M.S., 1996,”Cablebolting in

Underground Mines”, BiTech Publishers Ltd., Manitoba

Sulaiman, M.S.,2007, “Analysis of Stope Stability by Considering the

Discontinuities Resulted from Observation Activity at South

Ciurug Mine Area, Pongkor Underground Gold Mine, PT. Antam

Tbk.”, Magister Thesis (Indonesian), Mining Engineering Post

Graduate Program, Fac. of Mining and Petroleum Engineering,

ITB.

Sulistianto, B., Rai, M.A., Kramadibrata, S., Hartami, P.N., Matsui, K.,

Nakagawa, H. and Setiawan, I.D.., 2003, “Determination of Insitu

Stress Using Hydraulic Fracturing Method at Pongkor

Underground Gold Mine, West Java, Indonesia”, Proc. of the 3rd

International Symposium on Rock Stress, 4-6 November 2003,

Kumamoto, Japan, Sugawara, Obara & Sato (eds), A.A. Balkema

Publ, pp 383-388.

Sulistianto, B., Wattimena, R.K., Kramadibrata, S., Sulaiman, M.S.,

Matsui, K., and Ardianto, A., 2007, “Borehole Investigation To

Grasp The Condition Of Stope’s Roof In Cut-and-Fill Mining At

Level 500 Ciurug, Pongkor Underground Gold Mine, Indonesia”,

International Workshop on Earth Science and Technology,

Fukuoka, Japan.

Documents

Determination of Stope Geometry in Jointed Rock Mass