Geological dynamic hazards during coal mining can be caused by the structural instability of a composite system of roof rock and coal layers, and joints in coal play a vital role in this structural instability. In this paper, uniaxial compression simulation tests on roof rock–coal combined bodies with a single joint at different angles in coal were conducted using PFC2D software. In particular, the effects of joint angle on the uniaxial compressive strength (UCS), acoustic emission (AE) and failure characteristics in a combined body were analysed. The results show the following. (1) The joint in coal causes a decrease in the UCS and crack initiation stress, and has promoting effects on the AE occurrence of the combined body. With an increase in the included angle α between the loading direction and joint plane direction, the UCS, AE occurrence time and crack initiation stress first decrease and then increase, presenting ‘V-shaped’ curves. The UCS and crack initiation stress at α of 45° are the lowest, and the AE occurrence time at α of 30° is the earliest. (2) The AE characteristics at α of 0°, 75° and 90° are basically consistent with those of the intact combined body, whereas at α between 15° and 60°, the intensity and frequency of AE events in the macro-failure stage are reduced by the joint in coal. (3) The failure of the combined body mainly occurs within the coal, whereas no apparent failure is observed for the roof rock. Only when α is between 15° and 45°, the roof rock near the bedding plane is destroyed by crack propagation in the coal. (4) Three typical failure modes (i.e. V-shaped shear failure cutting across the coal, shear failure partly along the joint plane direction and shear failure along the joint plane direction) are found for combined bodies with a single joint in coal.

In deep hard‐rock mines, the failure of subsurface structures (e.g., tunnels, stopes, and shafts) has been a significant problem affecting mining safety due to the high‐stress environment. In this paper, the mechanism of structural failure and instability is discussed, and optimized excavation methods are proposed for stress control in deep gold mines. Based on the field observation and investigation of the joints distribution and rock failure modes at 800–1200 m depth of several large gold mines and a typical ultradeep borehole (2017 m depth) in northwest Jiaodong Peninsula, three engineering methods for reducing stress, including the stress transferring by mining optimizations, pressure relief by boreholes, and energy release in advance by optimizations of excavation and support, are analyzed by numerical simulation and field monitoring. Results show that stress reduction by excavation alone is limited and the backfill mining method is more conducive to stress transfer than the opening stope method. Roof contacted backfill can produce an unloading zone around the stope and reduce the stress of the surrounding stope. Relief boreholes can reduce the stress concentration of stopes, but the effect of cutting seams generated by presplitting blasting on pressure relief is not significant. The technology “short excavation and short support” releases less energy. By increasing the bench height and the reasonable timing of support by calculating, the elastic strain energy of rock in the shaft is prereleased, which benefits the long‐term stability of the shaft.

In order to study the present state of the tectonic stress field in northwest Jiaodong Peninsula, the characteristics of the ground stress distribution along depth were analyzed with the method of regression analysis. A total of 164 items of in situ stress measurement data were collected. There are mainly two types of in situ stress states, one is σ_H>σ_h>σ_v and the other is σ_>σ_v >σ_h, and the type of in situ stress state is related to the depth. The principal stresses σ_H, σ_h, and σ_v increase approximately linearly with depth, and the stress gradients are 0.0556, 0.0248, and 0.0346, respectively. The lateral pressure coefficients KH, Kh, and Kav vary approximately hyperbolically with the increase in depth and approach 1.99, 1.12, and 1.56, respectively. The ratio of half of the maximum horizontal differential stress to the average horizontal principal stress μ_d varies linearly with the increase of depth and approaches the value of 0.33. The maximum shear stress τ_m increases approximately linearly with depth, and the stress gradients are 0.0146. The ratio of horizontal principal stress difference with vertical stress on 1000 m–2000 m is 0.5– 0.7. In addition, the maximum level of principal stress advantage direction in the northwest Jiaodong Peninsula is nearly E–W.

Increasing construction depth, changeable failure formations, and redistributed stress conditions inevitably make the mechanical response of mining shaft rock mass complicated. The design method and construction technology of deep shaft need to be perfected urgently. We studied the stability and control method of a main shaft from −930 m located in a deep gold mine. It is concluded that the surrounding rock of the shaft faces conditions including non-high-stressed, high-stressed, and possible instabilities. The failure types include structural plane-controlled failure, deep stress induced failure, rock burst, low confining pressure shear failure, spalling, and frequent conversion of multiple failure. The damages of the surrounding rock mass at −930~−1500 m displayed “ear-shaped” damage, with damage ranges of less than 2.5 m. The shaft temporary reinforcement adopted bolt mResh beam support, in which the length of the bolt was 2.5~3 m, and the row spacing was 1.5 m. The steel type and diameter were determined by the on-site bolt pull test. This temporary support countermeasure plays an important role in preventing shaft deformation and is worth promoting in similar mines.

A series of true triaxial compression (TTC) tests was conducted on granite specimens to investigate their fracture behavior. The granite was sampled from a 1500‐m‐deep exploration borehole located on the Sanshandao gold field in Shandong Province, China, where a 1915‐m deep vertical shaft was planned. Under true triaxial stress, the fracture of the granite specimen shows anisotropic characteristics, which is manifested as different strains in the three directions of the specimen, and the macroscopic fractures formed are parallel to the direction of maximum principal stress σ₁ and intermediate principal stress σ₂. In addition, the effects of the intermediate principal stress σ₂ and the minimum principal stress σ₃ on the fracture of granite specimen are different. A higher σ₂ causes the post‐peak curve of the granite specimen to develop toward a Class‐II type of curve, and the failure mode develops toward splitting failure, while a higher σ₃ causes the post‐peak curve of the granite specimen to develop toward a Class‐I type of curve, and the failure mode develops toward shear failure. In particular, under the stress regime whereby σ₃ = 0, σ₂ = 40 MPa, it simulates the tangential stress adjustment process when the shaft is excavated to a depth of 1000 m. The fracture mode of the granite specimen and the post‐peak curve characteristics both indicate that the shaft excavation to this depth has the risk of inducing a strain burst.