Field investigations have proven that layer‐crack structures make an essential contribution to rock burst occurrence in brittle coal or rock mass. This paper first studies the mechanical behavior of layer‐crack red sandstone specimens with different geometric configurations of vertical fissures (ie, fissure width, fissure length, and fissure number) through uniaxial compression and Brazilian tests. Then, the digital speckle correlation method (DSCM) and acoustic emission (AE) technique are used to record the deformation and failure processes. Finally, the failure mechanisms of the layer‐crack structure are presented and discussed. Under compressive conditions, the failure process can be divided into compatible support, incompatible support, and postpeak unstable failure stages. An increase in fissure width, as well as fissure length or number, causes an increase in the inner damage accumulation and a decrease in the stiffness coefficient by decreasing the strain of the compatible support stage. Thus, the bearing capacity of the layer‐crack structure is weakened. Under tensile conditions, the tensile ability decreases linearly as the fissure length increases, but the vertical pressure acting on the fissure increases linearly as the fissure width increases. Thus, the tensile strength of the layer‐crack structure decreases as the fissure length (or width) increases.

Waste-rock material used in underground backfill mining has a granular texture and acquires non-linear deformation characteristics when compressed. The deformation modulus of waste-rock measured by a laboratory compression test is significantly different from the true deformation modulus in the field, due to the complete confining effect of the loading steel cylinder. In this study, we performed a series of laboratory-based compression tests on waste-rock samples. The results showed that lab-acquired deformation modulus variations of waste rock could be divided into three stages: slow increase, accelerated increase, and rapid increase. We also measured the true deformation modulus of backfill waste rock by conducting a field test in gob areas of the Tangshan coal mine, China. The hardening process of backfill waste rock during the field test was analyzed, and could be divided into four stages: roof contact, rapid compression, slow compression, and long-term stable. With the increase of axial strain, the lab- and field-measured deformation moduli of waste rock both increased exponentially. A correction parameter was proposed to investigate the relationship between the field-generated true deformation modulus and the lab-tested deformation modulus. The correction parameter k positively correlated with the axial strain, in the form of an exponential function. The magnitude of k was between 0.5616 and 0.6531.

Destress drilling method is one of the commonly used methods for mitigating rock bursts, especially in coal mining. To better understand the influences of drilling arrangements on the destress effect is beneficial for rock burst mitigation. This study first introduced the rock burst mitigation mechanism of the destress drilling method and then numerically investigated the influences of drilling arrangements on the mechanical properties of coal models through uniaxial compression tests. Based on the test results, the energy evolution (i.e., the energy dissipation and bursting energy indexes) influenced by different drilling arrangements was analyzed. When the drilling diameter, the number of drilling holes in one row, or the number of drilling rows increases, the bearing capacity of specimens nonlinearly decreases, but the energy dissipation index increases. In addition, the drilling diameter or the number of drilling holes in one row affects the failure mode weakly, which is different from that of the number of drilling rows. Consequently, the bursting energy index decreases as increasing the drilling diameter or the number of drilling holes in one row, but as increasing the number of drilling rows, the variation law of bursting energy index is not obvious. At last, the influencing mechanism of drilling arrangement on the rock burst prevention mechanism of the destress drilling method was discussed and revealed.

Aiming at the problems of large deformation, long duration, and difficult control of floor heave in gob-side entry driving in weakly cemented soft rock, this paper takes the weakly cemented soft rock mining area in Western China as the engineering background, and studies the characteristics and mechanism of floor heave in gob-side entry driving in weakly cemented soft rock by means of a field investigation, physical component analysis, mechanical property tests of the surrounding rock, and the stress monitoring of the surrounding rock. The classification control method of floor heave is put forward, and field tests are conducted. The results show that: (1) The floor heave characteristics of the dynamic change in the floor heave peak position of gob-side entry driving from the coal pillar side to the mining side are obtained through field observation. (2) Based on the analysis of field data and laboratory test data, it is concluded that the stability time of the overlying strata in gob-side entry driving is about 8 to 12 months. The main internal cause of roadway floor heave is the low load resistance of weakly cemented soft rock. High stress and strong disturbances are the main power sources of strong floor heave. The mechanism of floor heave affected by dynamic lateral abutment pressure is summarized, and the classification control method of floor heave is proposed. (3) The gob-side entry driving support technologies of “adjusting excavation deployment” and “surrounding rock pressure relief and improving support” are proposed. Through field tests, the floor heave can be effectively controlled.

During coal mining, the deformation and failure of a weakly cemented soft rock roadway roof could cause difficulties for roadway support. In this paper, a combination of on-site measurement and theoretical analysis is used to solve this issue. Firstly this paper investigates the in situ deformation and failure behaviors of a soft rock roadway in a mine in Western China. Then, the failure mechanism and corresponding support principles are discussed and given. Third, various support schemes (bolt and cable reinforcement optimization, grouting, and single prop + top beam combined reinforcement) are proposed and tested. Results show the support capacity can meet the requirements after optimizing the bolt and cable reinforcement support. Due to the development of roof cracks and low grouting pressure, the grouting slurry did not completely fill the roof cracks, resulting in a poor roof control effect. The passive support of a “single prop + top beam” can effectively control the roof subsidence and achieve good application results.