CIM Bulletin, Vol. 2, No. 6, 2007
M.R. Saharan, A.K. Chakraborty, A. Sinha, N.K. Babar, H.R. Kalihari, and C.P.N. Pathak
The orebody in Chikla Mine, particularly between the -70 and -170 ft levels, displays two systematic joint sets: one has an average spacing of 2.5 m, with its strike direction along N60°E and dipping almost vertical due east; and the other has an average spacing of 2.0 m, with its strike direction along N80°W and dipping almost vertical due south. Schistocity planes, which are horizontal and sub-horizontal, form the third plane of weakness in the orebody. Such planes exhibit a spacing varying from 5 to 30 cm. All these weakness planes are devoid of any infillings, have rough, planar, and wavy surfaces, and are very tight in nature. The orebody has an average uniaxial compressive strength (UCS) value of about 120 MPa, as found from rebound values of a standard Schmidt Hammer. Results of preliminary rock mass characterization tests indicate that the orebody has Barton’s Q value ranging from 16 to 38 (good rock mass), while Bieniawski’s RMR values range from 80 to 87 (very good rock mass). The host rock has been determined to be weaker than the orebody with Q values ranging from 5 to 8 (fair rock mass) and RMR values ranging from 49 to 63 (fair rock mass). Based on these classifications, it is observed that the orebody of Chikla falls under the engineering categories of “no supports require” and “spot bolting.” Numerical modelling results are in line with the results of rock mass classification, which indicate skin failures at roof level for the cut-and-fill stopes, taking into consideration the lower bound RMR values. Further, the aforementioned results have also been compared with field observations. The mine has already completed an experimental stope design and construction, using an open stoping operation. The roof (crown pillar) of this stope has been standing without any sign of failure for the past six months (i.e., six months subsequent to the completion of the stopping). Also, the measurements from hanging cables in the cutand-fill stopes further indicate that the cables are not carrying a load over 5 t.
It is concluded that the practice of central-grouted rock bolting may safely be discontinued as it appears to be an overly conservative design. Moreover, the grid of cable bolting may be safely increased from 2.0 m x 2.0 m to 2.5 m x 2.5 m before adopting the category of “no supports require.” It has been suggested to mine management that the above results should be verified from a rock mechanics instrumentation program covering instrumented cable bolts, stress metres, and strain bars in an experimental stope, prior to the adoption of the “no supports require” category.
Results of scientific investigations on the optimization of the rock reinforcement system (cable bolting) at Chikla Mine, MOIL, India, are presented in this paper. The mine is implementing a systematic roof support program that primarily contains cable bolting in a 2.0 m square grid pattern. The fully grouted cables are 16 mm in diameter and 12 m in length. Additionally, 20 mm diameter, 1.5 m long, full column grouted rock bolts in the centre of the cable bolting grid are also implemented for the manganese orebody with widths ranging from 9 to 24 m (average width 12 m) and a dip angle of 55° to 90°. Manganese ore mineralization, primarily in the form of braunite and gondite minerals, is flanked by muscovite schist/quartz muscovite, and schist in Chikla Mine, MOIL. The depth of workings is in between 40 and 70 m from the surface, and a cut-and-fill stoping operation is being practiced to produce average stope widths of 12 m. Field and laboratory investigations are carried out to obtain empirical rock mass classification parameters and input parameters for numerical modelling. The safety factor contouring method, using a failure criterion suggested by Sheorey (1997), is used for plain strain numerical modelling in a general purpose finite difference method-based numerical modelling code - FLAC3D.
