The protective layer mining method of the traditional deep coal seam in has been confronted with great challenges, and it is difficult for coal and gas to be extracted together. Taking the occurrence conditions of III1 mining area of Luling Coal Mine located at Huaibei, China, as engineering background, the influence law of the lithology on stress environment in front of the stope was analyzed by theoretical analysis and numerical simulations. The mining‐induced mechanical effect of coal‐rock mass was studied under different protective layer mining modes. The results showed that the peak value of the advanced abutment pressure decreased with the decrease of lithologic strength under the same mining conditions. For simulated geological conditions, the stress concentration coefficient of soft rock and coal seam protective layer mining modes was 1.9 and 1.7, respectively. Under the mining stress path of different protective layers, the ratio of axial stress increase and confining pressure unloading in secondary unloading phase were 2 : 1 and 1.5 : 1, respectively. The axial stress‐strain curves of different protective layer mining modes had similar trends, and they had a volume expansion at the end of unloading (failure stage). In addition, it revealed the pressure‐relief antireflection mechanism of the protective layer mining. Under the same confining pressure condition, the peak stress and peak strain increased with the increase of loading and unloading velocity ratio. The reduced value of the confining pressure increased, while the volume expansion decreased at failure. The results were applied to III1 mining area in soft rock protective layer mining, which created the mining way of traditional coal seam protective layer. Furthermore, the gas control technology of soft rock protective layer working face was put forward for deep coal seam with low permeability and high gas, enriching the pressure‐relief mining theory of protective layer.

In order to effectively address stability control technology issues of soft surrounding rock roadways in deep mines. This study analyses the deformation and failure characteristics of the surrounding rock of a −962 m horizontal track roadway with original support conditions based on a severe deformation case that occurred in a mine. Upon establishing a mechanical model of surrounding rock failure zoning for circular roadways, which is based on the relationship between the stress–strain curve of soft rocks and the secondary stress distribution and strength of surrounding rock, this study explores the influence of rock strength indicators, disturbance degree, and support resistance on the stress distribution of the surrounding rock. The failure or instability mechanism of high-stress soft and weak surrounding rock is revealed on this basis. A multi-stage strengthening combined support technology is proposed, which consists of “high-strength prestressed anchor bolt (cable) supports as the core, deep and shallow hole groutings as the foundation, bottom angle, and floor anchorage grouting reinforcements as the key.” Moreover, numerical simulation and engineering practice optimize and verify the support scheme. The results show that after adopting the multi-stage strengthening combined support technology, the deformation of the surrounding rock of the roadways was only 12.6~14.3% of that under the original supporting parameters, and the deformation rate was still less than 0.2 mm/d even after 40 days. The proposed surrounding rock support method realizes the stability control of the roadway, which also has specific reference significance for similar projects.

To explore the broken energy change and the specimen fragment influence of granite where the length-to-diameter ratio is 0.52, the SHPB device was used to perform dynamic loading on the granite specimens. The rock energy evolution law was analyzed by the energy time history curve, and according to rock fragment characteristics, the rock fractal dimension was calculated. The experimental results show that the rock energy-time history curve can be divided into four stages. The incident energy is independent of the length-to-diameter ratio of the rock specimens. When the length-to-diameter ratio of the rock specimens is 0.50.9, the difference of incident energy, transmitted energy, and reflection energy of rock specimens is small. With the length-to-diameter ratio increasing, the rock fragment size became larger. These rock fragments have good self-similarity. The rock specimen fractal dimension is at least 1.94, and the maximum D is 2.536. And with the length-to-diameter ratios increasing, the fractal dimension of rock specimens decreases. With the specimen fractal dimension increasing, the energy dissipation density of rock specimens also increases. The higher the energy dissipation density of a rock specimen, the more uniform rock fragments are.

Due to the influence of deep high stress, geothermal heat, and other factors, the law of desorption of methane in coal seams is more complicated in the process of mining deep coal seams, which is prone to methane over‐limit, coal and gas outburst, and other accidents. In order to study the desorption characteristics of coalbed methane under different loading and temperature conditions, the desorption tests at different deformation stages of coal containing methane were carried out in the process of loading‐adsorption‐desorption‐reloading until the coal sample was destroyed by using the seepage‐adsorption‐desorption test system on coal and rock mass, and the test programs were different combinations of gas pressure 1.2 MPa, two kinds of confining pressure, and three kinds of temperature. The results show that the cumulative methane desorption amount corresponding to each deformation stage presents a convex parabolic increase trend with the increase in desorption time, while the desorption rate presents a power function decay trend. Under the condition of the same desorption time, the cumulative methane desorption amount from large to small is residual deformation stage, compaction stage, near the peak stress, plastic deformation stage, and elastic deformation stage. Under the same confining pressure, temperature, and methane pressure, the maximum desorption rate from large to small is residual deformation stage, near the peak stress, plastic deformation stage, compaction stage, and elastic deformation stage. The desorption and diffusion of methane are promoted under the higher temperature and lower confining pressure, which presents a certain mechanism of promoting desorption. The thermal movement of methane molecules is intensified with the increase in temperature, and the adsorption effect between methane molecules and the molecules at the surface of the coal is weakened. The cumulative methane desorption amount and the maximum desorption rate increase with the increase in temperature. The cumulative methane desorption in the residual deformation stage is obviously greater than that in other deformation stages. The increase in confining pressure inhibits the development and expansion of pore fractures in raw coal specimen and hinders the increase in the effective desorption surface area. The cumulative methane desorption amount and the maximum desorption rate decrease with the increase in confining pressure.

In order to investigate the dynamic mechanical properties of the fractured rocks reinforced with different slurries, the prefractured and grouted sandy mudstone specimens were tested on a Split Hopkinson Pressure Bar (SHPB), and their dynamic failure processes were shot via a high‐speed camera. The test results showed that under dynamic loading, grouting reinforcement can relieve the rock failure at the fracture part. With the increase of the impact speed, the dynamic specific strength of the modified epoxy resin grouting increases from 0.85 to 1.01, and its specific strain rate increases from 1.09 to 1.04; the dynamic specific strength and specific strain rate of cement slurry grouting increase from 1.78 to 0.95 and from 0.60 to 0.98, respectively. As a whole, the specimen reaches or approaches the dynamic parameters of the intact rock specimen after grouting repair, the stability of the rock reinforced by modified epoxy resin grouting is superior to that of the rock reinforced by cement slurry grouting, and the latter can acquire high dynamic strength under low‐speed impact. The improved damage‐type Zhuwangtang (ZWT) model can be used to describe the dynamic mechanical behaviors of the fractured rocks reinforced by grouting very well.