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The use is discussed of an artificial roof for back-fill mining involving deep-hole caving in soft and fractured rock. Numerical models of the stope were developed using the FLAC-3D software and the effects of roof thickness on stope stability, stress distribution and roof settlement were determined. The results indicated that the plastic zones, stress-affected areas and the settlement all decreased with increasing roof thickness, but the extent of the decrease became lower with increasing thickness of the artificial roof. Using settlement as a criterion, the results showed that at a thickness of 9 m the maximum settlement would be 7.67 cm, which would meet mine safety requirements., The use is discussed of an artificial roof for back-fill mining involving deep-hole caving in soft and fractured rock. Numerical models of the stope were developed using the FLAC-3D software and the effects of roof thickness on stope stability, stress distribution and roof settlement were determined. The results indicated that the plastic zones, stress-affected areas and the settlement all decreased with increasing roof thickness, but the extent of the decrease became lower with increasing thickness of the artificial roof. Using settlement as a criterion, the results showed that at a thickness of 9 m the maximum settlement would be 7.67 cm, which would meet mine safety requirements.

Development of an evaluation index system of rock burst using six quantitative indicators - uniaxial compressive strength, tensile strength, tangential stress, rock-brittleness coefficient, stress coefficient and elastic-energy index - is described. Three numerical descriptors for each evaluation indicator corresponding to each class of rock burst intensity (non, light, medium or violent) are calculated using the backward cloud model, the weight of each index is calculated using the entropy method and the forward cloud model is established. The performance of the entropy-cloud model is determined using data from 209 sets of rock-burst examples from over 15 global underground-engineering projects, the sensitivity of the indicators is analysed, and the class of rock burst is discriminated. The results indicate that the sensitivity of the indicators decreases in the order, stress coefficient, tangential stress, elastic-energy index, rock-brittleness coefficient, tensile strength and uniaxial compressive strength, the accuracies of the 42 prediction sets using the entropy-cloud, Bayes, k-nearest neighbours algorithm and random forest models are 76.2, 67, 52, and 73.8%, respectively, and the accuracies of the 167 training sets are 82, 78, 60 and 86%, respectively., Development of an evaluation index system of rock burst using six quantitative indicators - uniaxial compressive strength, tensile strength, tangential stress, rock-brittleness coefficient, stress coefficient and elastic-energy index - is described. Three numerical descriptors for each evaluation indicator corresponding to each class of rock burst intensity (non, light, medium or violent) are calculated using the backward cloud model, the weight of each index is calculated using the entropy method and the forward cloud model is established. The performance of the entropy-cloud model is determined using data from 209 sets of rock-burst examples from over 15 global underground-engineering projects, the sensitivity of the indicators is analysed, and the class of rock burst is discriminated. The results indicate that the sensitivity of the indicators decreases in the order, stress coefficient, tangential stress, elastic-energy index, rock-brittleness coefficient, tensile strength and uniaxial compressive strength, the accuracies of the 42 prediction sets using the entropy-cloud, Bayes, k-nearest neighbours algorithm and random forest models are 76.2, 67, 52, and 73.8%, respectively, and the accuracies of the 167 training sets are 82, 78, 60 and 86%, respectively.

A discussion of stability factors in an undercut space in a fragment orebody including wedge geometry (size, shape and spatial location), undercut geometry (trend, plunge and cross-section), joint properties and field stress (stress orientation, field stress value and cohesion friction angle) is presented. For the undercut space with different trends, plunges and cross-sections, orthogonal simulation tests with three influence factors and five levels are carried out using Unwedge software. The results indicate that the base value of the wedge safety factor is assigned as 1.2 and the optimal parameters in the undercut space are trend of 315-325 degrees, plunge of 5 degrees and cross-section of 27 degrees (safety factor greater than 1.872). An in-situ test conducted at Kalatongke underground copper-nickel mine, Fuyun County, Xinjiang, Junggar Basin, China to verify the optimal parameters, using a trend value of 320 degrees, shows that the mine discharging is continuous in the undercut with no large-scale blockages., A discussion of stability factors in an undercut space in a fragment orebody including wedge geometry (size, shape and spatial location), undercut geometry (trend, plunge and cross-section), joint properties and field stress (stress orientation, field stress value and cohesion friction angle) is presented. For the undercut space with different trends, plunges and cross-sections, orthogonal simulation tests with three influence factors and five levels are carried out using Unwedge software. The results indicate that the base value of the wedge safety factor is assigned as 1.2 and the optimal parameters in the undercut space are trend of 315-325 degrees, plunge of 5 degrees and cross-section of 27 degrees (safety factor greater than 1.872). An in-situ test conducted at Kalatongke underground copper-nickel mine, Fuyun County, Xinjiang, Junggar Basin, China to verify the optimal parameters, using a trend value of 320 degrees, shows that the mine discharging is continuous in the undercut with no large-scale blockages.

The basic principles of genetic algorithms are expounded and an algorithm for optimising the structural parameters of stopes is introduced. It is shown by the computation of individual cases that reliable structural parameters can be obtained using the algorithm, which does not have the defects of conventional algorithms., The basic principles of genetic algorithms are expounded and an algorithm for optimising the structural parameters of stopes is introduced. It is shown by the computation of individual cases that reliable structural parameters can be obtained using the algorithm, which does not have the defects of conventional algorithms.

The discrete element method was used to calculate the stability of three types of goaf in a fractured ore section at the Gaofeng mine in China, taking into account differences in goaf size, the surrounding conditions and the initial rock stress field. The results showed that in goafs completely embedded in cataclastic ore sections, the plastic zone was completely distributed in the left section of the goaf roof, with no plastic zone occurring in the pillar. The small goaf was stable, with only occasional occurrences of falling of small blocks. The plastic zone in the roof of the medium size goaf was 7 m with a span of more than half of the roof length and was unconnected, so that a large-scale caving accident could not occur. The plastic zone in the roof of the largest goaf was 20 m with a span of more than half of the roof length. Since the plastic zones are interconnected a large-scale caving accident was inevitable., The discrete element method was used to calculate the stability of three types of goaf in a fractured ore section at the Gaofeng mine in China, taking into account differences in goaf size, the surrounding conditions and the initial rock stress field. The results showed that in goafs completely embedded in cataclastic ore sections, the plastic zone was completely distributed in the left section of the goaf roof, with no plastic zone occurring in the pillar. The small goaf was stable, with only occasional occurrences of falling of small blocks. The plastic zone in the roof of the medium size goaf was 7 m with a span of more than half of the roof length and was unconnected, so that a large-scale caving accident could not occur. The plastic zone in the roof of the largest goaf was 20 m with a span of more than half of the roof length. Since the plastic zones are interconnected a large-scale caving accident was inevitable.

The vertical distribution characteristics of Pb, Zn, Cu and As in a tailings deposit were investigated and the pollution and potential degree of ecological hazard assessed. It was found that the heavy metal concentration in the soil was lower than at the soil-tailings interface, where the pollution was strongest, but still higher than the soil background values; pollution within the tailings and at depth was also severe. Cu was less polluting than the other metals while the degrees of Pb, Zn and As pollution were different at different depths. The greatest environmental threat was posed by As, followed by Pb. It is concluded that a simple soil cover is insufficient to reduce the risks from heavy metals in tailings and that it needs to be combined with other effective measures., The vertical distribution characteristics of Pb, Zn, Cu and As in a tailings deposit were investigated and the pollution and potential degree of ecological hazard assessed. It was found that the heavy metal concentration in the soil was lower than at the soil-tailings interface, where the pollution was strongest, but still higher than the soil background values; pollution within the tailings and at depth was also severe. Cu was less polluting than the other metals while the degrees of Pb, Zn and As pollution were different at different depths. The greatest environmental threat was posed by As, followed by Pb. It is concluded that a simple soil cover is insufficient to reduce the risks from heavy metals in tailings and that it needs to be combined with other effective measures.

The Kelatongke Cu-Ni mine in Xinjiang, China, employs blasthole stoping with backfilling. Field tests were carried out during the construction of the second layer of the artificial roof involving the installation of strain gauges and joint meters after extraction of the initial drifts and backfilling had been completed. The stress and displacement of the backfill were monitored as mining and backfilling were progressing in other drifts. The results showed that the compressive stress of the backfill was gradually transformed into tensile stress and the displacement increased with time up to a maximum of 22.4 mm. If backfilling is not carried out immediately after extraction shear stresses may be induced in the surrounding drifts., The Kelatongke Cu-Ni mine in Xinjiang, China, employs blasthole stoping with backfilling. Field tests were carried out during the construction of the second layer of the artificial roof involving the installation of strain gauges and joint meters after extraction of the initial drifts and backfilling had been completed. The stress and displacement of the backfill were monitored as mining and backfilling were progressing in other drifts. The results showed that the compressive stress of the backfill was gradually transformed into tensile stress and the displacement increased with time up to a maximum of 22.4 mm. If backfilling is not carried out immediately after extraction shear stresses may be induced in the surrounding drifts.

The Phase2D software was used to simulate the mining-induced stress between ore sections after backfilling the goafs in an extended orebody during overhand and underhand cut-and-fill mining. For underhand mining, the calculated maximum compressive stress of 18 MPa was lower than the compressive strength and the maximum tensile stress of 0.47 MPa was much lower than the tensile strength value of 1.2 MPa for the roof filling body. The results indicate a safety factor coefficient of up to 1.57 for the filling body roof. For overhand mining, the predominant tensile stress had a maximum value of 0.79 which was lower than the tensile strength of the roof and higher than the shear strength of a weak structural plane in the roof, resulting in roof instability due to the destruction of the weak structural plane under the impact of the tensile stress. The results indicated that the tensile stress developed during overhand mining would result in roof falls and unsafe mining operations in the stope, and that the underhand mining method should be employed to optimise mining-induced stress., The Phase2D software was used to simulate the mining-induced stress between ore sections after backfilling the goafs in an extended orebody during overhand and underhand cut-and-fill mining. For underhand mining, the calculated maximum compressive stress of 18 MPa was lower than the compressive strength and the maximum tensile stress of 0.47 MPa was much lower than the tensile strength value of 1.2 MPa for the roof filling body. The results indicate a safety factor coefficient of up to 1.57 for the filling body roof. For overhand mining, the predominant tensile stress had a maximum value of 0.79 which was lower than the tensile strength of the roof and higher than the shear strength of a weak structural plane in the roof, resulting in roof instability due to the destruction of the weak structural plane under the impact of the tensile stress. The results indicated that the tensile stress developed during overhand mining would result in roof falls and unsafe mining operations in the stope, and that the underhand mining method should be employed to optimise mining-induced stress.

Numerical simulation with the LS-DYNA software was used to investigate the effect of dynamic loading resulting from underground medium-length hole blasting on the stability of the backfill. The feasibility of constructing an artificial pillar using waste rock grouting was also investigated. The results showed that a pre-splitting crack could effectively reduce the impact of blast dynamic loading and the effect increased with increasing length of the pre-splitting crack. At a pre- splitting crack length of twice the width of the ore break resulting from blasting the vibration peak decreased by 83.07% and the maximum tensile stress by 91.14%, which corresponded to the required backfill strength. The main vibration frequency decreased and the decay period increased in the presence of the pre-splitting crack. The vibration reduction effect decreased with an increase in millisecond delay time. The results also showed that medium-length hole blasting could be carried out using an artificial pillar made of rubble backfill provided that reasonable blast parameters were employed., Numerical simulation with the LS-DYNA software was used to investigate the effect of dynamic loading resulting from underground medium-length hole blasting on the stability of the backfill. The feasibility of constructing an artificial pillar using waste rock grouting was also investigated. The results showed that a pre-splitting crack could effectively reduce the impact of blast dynamic loading and the effect increased with increasing length of the pre-splitting crack. At a pre- splitting crack length of twice the width of the ore break resulting from blasting the vibration peak decreased by 83.07% and the maximum tensile stress by 91.14%, which corresponded to the required backfill strength. The main vibration frequency decreased and the decay period increased in the presence of the pre-splitting crack. The vibration reduction effect decreased with an increase in millisecond delay time. The results also showed that medium-length hole blasting could be carried out using an artificial pillar made of rubble backfill provided that reasonable blast parameters were employed.

Blasting crater tests were carried out using 12 single holes and 4 variable distance holes in a tunnel with similar rock properties as a mining stope. The characteristic blasting crater curves for the emulsion explosives used, the relationships between blasting crater volume and charge depth and between crater radius and charge depth, the optimum and critical depths for the blasting crater and the blast fragmentation were investigated. The results showed that the optimum and critical depths for the blasting crater were 1.09 and 1. 65 m, respectively, the optimum blasting crater radius was 0.79 m, the optimum blasting crater volume was 0.37 m3, the hole spacing should be 1.9 times the optimum blasting crater radius, with an average 58.4% blast fragmentation., Blasting crater tests were carried out using 12 single holes and 4 variable distance holes in a tunnel with similar rock properties as a mining stope. The characteristic blasting crater curves for the emulsion explosives used, the relationships between blasting crater volume and charge depth and between crater radius and charge depth, the optimum and critical depths for the blasting crater and the blast fragmentation were investigated. The results showed that the optimum and critical depths for the blasting crater were 1.09 and 1. 65 m, respectively, the optimum blasting crater radius was 0.79 m, the optimum blasting crater volume was 0.37 m3, the hole spacing should be 1.9 times the optimum blasting crater radius, with an average 58.4% blast fragmentation.