Determination of stope geometry in jointed rock mass at Pongkor Underground Gold Mine

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1 Volume 5, Number 2, April 2009, pp [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., Jakarta530, Indonesia **** Member of ISRM: Dept. of Earth Resources Eng., Kyushu University, Fukuoka , Japan Received ; accepted 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 N330 0 E/ 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- 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 stopeback 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. JCRM All rights reserved.

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 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.

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.

Two bore holes were drilled in each crosscut with different orientation as can be seen in Figure 2 and Table 1. 2.2. Laboratory Testing Rock sample specimens obtained from the drilling were tested in laboratory.

2 64 B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp 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 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-. Two bore holes were drilled in each crosscut with different orientation as can be seen in Figure 2 and Table 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 Orientation DH Location Trend (N.. E) Plunge (.. ) Remarks Depth (m) GTC 1 Xcut Horizontal GTC 2 Xcut Inclined GTC 3 Xcut Horizontal GTC 4 Xcut Inclined Table 2. Physical and mechanical properties of rock No Material (gr/cm 3 ) E C (MPa) (.. 0 ) c (MPa) 1 Sidewall Andesite Quartz Vein Joint Note : : Rock Density C : Cohession E : Young s Modulus : Angle of internal friction : Poisson s Ratio σ c : 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- respectively. RQD values obtained from this drilling showed Figure 1. Research location and typical rock lithology Joint A Joint B Joint C Figure 2. Drilling at research location Figure 3. C Drill-core from Bore-hole GTC-02 at X cut-10

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.

Discontinuities Observation To observe the joint existence along the borehole, special equipment was used to grasp the joint position and the orientation.

3 B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp that the rock mass is in range of very poor to excellent condition. Figure 3 shows the example core condition obtained from this location 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 N110 0 E/47 0, N030 0 E/56 0 and N075 0 E/32 0 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-. Figure 5 shows the result of the observation plotted in stereonet. Since Xcut- located relatively in the upper part of Xcut-10 (figure 2), the joints showed in good persistence. 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-, 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- were 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- 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 Numerical method Figure 5. Discontinuity orientation at Xcut-10 (left) and Xcut- (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 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.

4 66 B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp 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/cm 3 ) E K G 1 Sidewall Andesite Quartz Vein C (MPa) (.. 0 ) Jkn (GPa/m) Jks (GPa/m) 3 Joint 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- (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 0kN 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 355mm 2 Meanwhile, at a stope-back of stope in front of Xcut-, 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 AS/NZS 1594 Gr.HA 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). N' Q' A B C...(1) RQD J Jn J a r Q'...(2) Where J n, J r and J a 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 Xcut-10 Xcut- Parameter Description Rating Description Rating RQD Three Sets + Three Sets + J n Random Random J r Rough, Undulating 3 Rough, Undulating 3 Zone of Zone of J a disintegrated disintegrated J w Minor inflow 1 Minor inflow 1 Low Stress, Near Low Stress, Near SRF Surface Surface Q RQD/J nxj r/j axj w/srf 0.43 RQD/J nxj r/j axj w/srf 0.4 Q A σ c/σ 1 = 32.59/4 0.8 σ c/σ 1 = 32.59/4 0.8 B α = α = C Stope back 2 Stope back 2 N

B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp.63-68 67 Figure 7.

height of opening was 4-5m considering the height of Jumbodrill, the hydraulic radius of stope-back was then calculated based upon the simulated length of unsupported excavation from 4m to 100m and

5 B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp Figure 7. Numerical model result of stope in x-cut-10 and x-cut- by using 1m 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 Jumbodrill, 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 Figure 8. Vertical deformation at stope-back of x-cut- by using 0.75m bolt spacing 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

75m if mining operation in the level of Xcut- is started. opening is in stable condition.

6 Stope Width (m) Span Widthout Support (m) Hydraulic Radius (m) 68 B. SULISTIANTO et al. / International Journal of the JCRM vol.5 (2009) pp 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- is started. 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 CONCLUSION Figure 9. Stability chart of research area Table 7. Stope condition related to opening geometry Remarks Xcut 10 Xcut Stable Stable Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone Transition Zone 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 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-, 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 3 rd International Symposium on Rock Stress, 4-6 November 2003, Kumamoto, Japan, Sugawara, Obara & Sato (eds), A.A. Balkema Publ, pp 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.

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