A shear strength reduction method based on a dual factoring strategy was considered in which cohesion and the coefficient of internal friction were each reduced by a different factor. By using a strength reduction path introduced in the present study, series of numerical tests were conducted to bring the slope into different states of critical equilibrium through which definitions of the factor of safety and the strength reduction path adopted in some preceding studies were evaluated. This strategy was found to be more conservative than the traditional shear strength reduction method. Generally, the factor of safety defined in the limit equilibrium method should be adopted, and the most probable strength reduction path under the principle of the shortest path could be an effective approach to obtain the minimum value of the factor of safety directly. To implement this strategy, a mathematical model based on limit analysis solutions is proposed to describe the shear strength components of a slope under different critical equilibrium states, and a general computational procedure is outlined.

The stability of the filling roof—as an important bearing unit in the stope of the access stope in the filling mining method—is of great significance to guarantee the safe and efficient production of the mine. Arrangement of the mining approach in downward cemented filling stope is the key factor affecting the stability of the filling body roof. Based on a combination of laboratory tests, theoretical analysis, and numerical simulation, the influence of different mining approach arrangements on the stability of the filling body roof is analyzed. The weak filling surface is formed between adjacent mining paths. The mechanical strength of the weak filling surface is significantly reduced by laboratory experiments. The relationship between the distribution of the weak filling surface, azimuth angle, and the stability of the filling roof is further studied by numerical simulation. The results show that, when the upper and lower layered mining approaches are arranged vertically or nearly vertically, the areas of stress concentration and the plastic zone in the numerical simulation results are the smallest, and the stability of the filler roof is the best.

The stability of the filling roof as an important bearing unit in the stope of the access stope in the filling mining method is of great significance to guarantee the safe and efficient production of the mine. Arrangement of mining approach in downward cemented filling stope is the key factor af-fecting stability of filling body roof. Based on combination of laboratory test, theoretical analysis and numerical simulation, the influence of different mining approach arrangement on stability of filling body roof is analyzed. The weak filling surface is formed between adjacent mining paths. The mechanical strength of the weak filling surface is significantly reduced by laboratory experiments. The relationship between distribution of the weak filling surface, azimuth angle and stability of filling roof is further studied by numerical simulation. The results show that the stability of the filler roof is the best when the upper and lower stratified mining approach is arranged vertically.

When jointed rock masses are in a high-stress environment, the roughness of the joints is the key factor controlling their shear strength. Their loading behavior is also different from the constant normal load (CNL) conditions controlled in conventional laboratories; rather, they follow the constant normal stiffness (CNS) conditions. To investigate the effects of normal stiffness and roughness on the shear mechanical properties of unfilled joint surfaces, shear tests were simulated using PFC3D (5.0) software under CNS conditions. The effects of normal stiffness of 0 (constant normal stress of 4 MPa), 0.028 GPa/m (low normal stiffness), 0.28 GPa/m (medium normal stiffness), and 2.8 GPa/m (high normal stiffness), and joint roughness coefficients (JRC) of 2~4 (low roughness), 10~12 (medium roughness), and 18~20 (high roughness) on the shear stress, normal stress, normal deformation, surface resistance index, and block failure characteristics of the joint surface were obtained. The results indicate that for different combinations of normal stiffness—JRC—the shear simulation process primarily exhibits three deformation stages: linear stage, yield stage, and post-peak stage. Shear stress increases initially and then decreases as shear displacement increases. When normal stiffness is no less than 0.28 GPa/m, both normal stress and JRC increase gradually with increasing JRC and normal stiffness. When the normal stiffness is no greater than 0.028 GPa/m, the normal stress shows no significant change. The normal displacement changes from “shear contraction” to “shear expansion” with increasing shear displacement and from positive to negative values while the displacement gradually increases; the maximum normal displacement decreases with increasing normal stiffness and increases with increasing JRC. The peak SRI value increases with increasing JRC and decreases with increasing normal stiffness. As normal stiffness increases, the number of tensile cracks for JRC 2~4 first decreases and then increases, while the number of shear cracks gradually increases; for JRC 10~12 and 18~20, both the number of shear cracks and tensile cracks increase with increasing normal stiffness. This paper simulates the actual mechanical environment of deep underground joints to expound the influence of normal stiffness and joint roughness on the stability of deep rock masses. The research results can provide certain theoretical references for predicting the stability of deep surrounding rocks and the stress of support structures.

Retaining wall is essential for stopes mining in two steps, for it can prevent the instability and collapse of backfill. In this study, taking the retaining wall of backfilled stope as the research object, a stability analysis method of retaining wall based on the close coupling of catastrophe theory and numerical analysis was proposed. First, by extracting the unit failure rate of the retaining wall from the numerical simulation results and fitting it with the mining depth, the functional expression between them was established. Second, the function relation was transformed into the normal form according to catastrophe theory, and the instability criterion of retaining wall was deduced. Furthermore, an effort was made to analyze the changing law of the state of retaining wall and calculate the critical span of stope, under different thickness conditions. On this basis, the application test of retaining wall was carried out by using this method. The results show that with the thickness decreasing, the values of splitting variables and show a reverse trend, which leads to the discriminant of instability criterion decreasing and turning from positive to negative, resulting in the collapse. Meanwhile, in order to ensure the stability, the wider the span of the stope, the thicker the retaining wall is required, and conversely, the thicker the retaining wall, the higher the adaptability to the span of stope. In addition, it can be found from the application test that instability was bound to occur with a thickness of 3 m, but the retaining wall with a thickness of 4 m maintained stable, which tended to be consistent with the analysis. Therefore, the stability analysis method proposed in this study provides a way to accurately evaluate the stability of the retaining wall and calculate the critical thickness of that, and its application value is expected to be further explored.

Parallelepipeds specimens were made to further investigate the rockburst occurrence mechanism of ore pillars in underground mining units. The investigation was carried out with uniaxial compression systems and real-time testing systems, such as stress, video, and acoustic emission, combined with digital image correlation (DIC) and SEM electron microscope scanning technology, to systematically analyze the evolution of rockburst of ore pillars, strain field characteristics, acoustic emission characteristics, mesoscopic characteristics of the rockburst fracture, morphology of the bursting crater, and debris characteristics. The findings demonstrate that the pillar’s rockburst process went through four stages, including the calm period, the particle ejection period, the block spalling period, and the full collapse period. According to DIC digital image correlation technology, the development of cracks in the rock is not obvious during the calm period, but during the small particle ejection and block spalling periods, the microcracks started to form and expand more quickly and eventually reached the critical surface of the rock, resulting in the formation of a complete macro-rockburst rupture zone. During stage I of the test, the rate of acoustic emission events and energy was relatively low; from stages II to IV, the rate gradually increased; and in stage V, the rate of acoustic emission events and energy reached its maximum value at the precise moment the rock exploded, releasing all of its stored energy. The specimen pit section primarily exhibits shear damage and the fracture exhibits shear fracture morphology, while the ejecta body primarily exhibits tensile damage and the fracture exhibits tensile fracture morphology. The location of the explosion pit is distributed on the left and right sides of the middle pillar of the specimen, and the shape is a deep “V”. The majority of the rockburst debris is greater than 5 mm, and it mostly takes the shape of thin plates, which is comparable to the field rockburst debris’s shape features.