Deep mining stress increase and thick hard rock layer endowment drive rock body fissure expansion, leading to sudden fracture of the roof plate accidents. The roof slab of 11,129 working face in Zhangji Mine is a 16 m thick sandstone composite roof slab, whose compressive strength reaches 89.8 MPa. The hardness of the roof slab is high, its bearing capacity is high, and it is difficult to collapse. The hardness of the roof plate is high, its bearing capacity is high, and it is difficult to fall. If the roof plate does not fall in time, it will lead to the energy accumulation of itself, and the sudden fracture and fall will make the working face and the two roadways show the drastic mine pressure, which will lead to the risk of the coal wall appearing the slice of gangs, the roof falling and the pressure frame. Therefore, for the 11,129 working face thick hard sandstone roof plate direct cover and mining technology conditions. It is very necessary and imperative to research the artificial roof release technology of deep hole pre-cracking blasting to improve the adventitious fall of the thick hard roof plate, reduce the overhanging roof area of the mining hollow area, and reduce the incoming pressure strength of the roof plate. This paper adopts the research method combining theory, physical experiment, numerical simulation and field test, focusing on research and analysis of the deflection change of the roof plate after pre-cracking and blasting, the relationship between the roof plate and the role of the bracket, and the effect of artificial blasting to release the roof. The results show that the theoretical part deduces and analyzes the relationship equation of the influence of the rotating and sinking of the roof plate on the pressure of the support, and combined with the law of the roof plate-support action can be divided into three stages: (1) Initial stage of mining, (2) Rotating and sinking stage, and (3) Breakdown and fragmentation stage. The deep stress increase and the rock layer's own characteristics (hardness and thickness) are the main controlling factors to determine the bearing capacity of the stent. As the depth of coal mining increases, the pressure at the minimum roof control distance of the stent gradually increases from 11.5 to 34.6 MPa. The increase of rock hardness relatively reduces the force acting on the stent, and the stent bearing capacity decreases by 75% when the hardness increases from 8 to 32 GPa. The 11,129 engineering geological parameters are brought into the calculation to obtain the pressure in the stent control distance of 13.5–20.5 MPa. The simulation part analyzes the stress relief degree of the roof plate before and after the roof blasting, and the peak stress in front of the coal wall is reduced from 46.7 to 30.9 MPa, which is an obvious effect of pressure relief, and the simulation results are analyzed in comparison with the theoretical results to validate the reliability of the simulation results. Further, the physical similarity experiments were conducted to analyze the deformation behavior of mining instability and crack evolution characteristics of the overburden rock of the quarry after cutting the roof, and the peak stress of the coal wall was in the range of 22.7–27.8 MPa after the blasting decompression, which was mutually verified with the theoretical analysis and the simulation results to ensure the reliability of the calculation results. Combined with the geological conditions of the mine area, the "fan-shaped" blasting hole arrangement is proposed to carry out segmental blasting, and the specific parameters of different locations of the holes are given. The effect of pre-cracking blasting was examined, and it was found that the rock mass at the hole opening had a high degree of fragmentation, and the ring and strike cracks appeared alternately in the middle and lower part of the drilling holes, and the polygonal energy slit played a very effective role, and achieved the purpose of artificially releasing the roof. At the same time, it is improved the load-bearing environment of the mining support, which further improves the safety and efficiency production of the coal mine.

The problem of strong ground pressure caused by thick and hard roofs in coal mine stopes has become more prominent. Blasting roof cutting and pressure relief are effective technical ways to solve this engineering problem. The key to achieving roof cutting and pressure relief is the evolution and penetration of cracks between blasting roof cutting holes in a mining stress environment. To solve this problem, research on the evolution of cracks between blasting roof cutting holes in a mining stress environment is being conducted. Along with specific engineering examples, the mechanical effect of blasting dynamic load in the surrounding rock medium between holes in a mining stress environment is analyzed using the research methods of field investigation, blasting simulation, and mechanical analysis. The fracture evolution law between blasting roof cutting holes under the conditions of different hole diameters and spacings, confining pressure environments, and rock mass strengths is analyzed using the univariate comparative analysis method. The research shows that the superposition effect of blasting dynamic load between adjacent blastholes in the transmission process leads to the “X” type penetration evolution of surrounding rock fracture area and fracture area between blastholes. Among them, the blasting effect of a hole diameter of 70 mm and a hole spacing of 0.8 m is best. When designing the blasting roof cutting scheme, we should pay attention to the key factors such as hole diameter and hole spacing. The hole spacing is particularly critical to the effect of roof cutting and crack formation, whereas the confining pressure and rock mass strength have little impact on the expansion and development of blasting cracks, which can be listed as secondary factors for comprehensive analysis in design. According to the transmission law of blasting dynamic load between holes, the critical criterion of blasting top clearance penetration under mining stress environment is deduced. Four important factors affecting blasting top cutting effect are simulated and studied for optimizing the design of roof cutting scheme. Practical engineering application results show that the blasting roof cutting scheme achieves a good seam- forming effect and creates good initial conditions for thick and hard roof cutting. It can serve as a reference for the decision making of blasting top caving technology under similar engineering conditions.

Aiming at the mining disaster of a thick hard roof, based on the analysis of the mining instability influence of the thick hard roof, this study constructs the mining bearing mechanical model of the thick hard roof by using mechanical theory and obtains the mechanical distribution equation of mining bearing and energy accumulation, the mining instability energy release equation, and the dynamic load response equation of a hydraulic support in the working face, as well as the dynamic load response characteristics of the hydraulic support in the working face, putting forward the technical countermeasures for the strong dynamic pressure control of the thick hard roof in the working face. This research shows that 1) the larger the overburden load and suspension span of the thick hard roof, the more serious the mining bearing state and energy accumulation evolution; the greater the rock thickness and elastic modulus of the thick hard roof, the greater the flexural stiffness of the roof, resulting in the increase of the roof mining limit breaking span, which indirectly aggravates the mining bearing state and self-energy accumulation evolution; 2) the dynamic support resistance of the hydraulic support is composed of the dynamic support resistance caused by the release of elastic energy accumulated by mining of the thick hard roof, the work done by the overlying load, and the static support resistance caused by the direct roof gravity; 3) the dynamic support resistance caused by the work of the overlying load accounts for the highest proportion, followed by the dynamic support resistance caused by the release of mining elastic energy by the thick hard roof; the cause of mining instability and the strong dynamic pressure of the thick hard roof lie in the large span of the mining suspended roof, and the large-scale mining suspension structure of the thick hard roof leads to a high overlying load and large accumulated energy; and 4) the mining instability of the thick hard roof leads to a strong dynamic load response of the hydraulic support; adopting pre-splitting and roof cutting technology to reduce the breaking span of the thick hard roof and reducing the impact dynamic load caused by mining instability of the thick hard roof can effectively eliminate the potential safety hazard of overlimit bearing of the hydraulic support.

Due to the change of ground stress environment caused by underground coal mining, the intense lateral mining stress concentration is formed around the stope; so section coal pillar is generally set up to bear the mining pressure, but the different sizes of coal pillars have obvious influence on the bearing capacity of those pillars and the characteristics of mining pressure. Mastering the mechanism characteristics by which coal pillars bearing capacity and mining stress distribution is crucial to identify the reasonable coal pillar size and give full play to the bearing role of section coal pillar, given their importance for the safety and bearing stability of engineering rock mass in underground coal mining. Therefore, the bearing characteristics of section coal pillar and the redistribution of mining stress are achieved with a mechanical model analysis on the basis of the analysis of coal pillar bearing and mining influence characteristics. Moreover, applying the elastic‐plastic mechanics theory revealed the mechanical equations of the effective bearing size of coal pillar and redistribution of mining stress in longwall face. Combined with the analysis of a specific engineering example, the research results are as follows. During a roadway excavation, the continuous mining stress transfer occurs “stress redistribution” and the mechanical failure of bearing coal pillar consists of lateral mining and roadway side failures. The bearing coal pillar has two critical dimensions (i.e., the critical dimension W_e of the self‐bearing stability coal pillar and the critical dimension W_p of failure through the coal pillar). The mechanical state of the lateral mining stress redistribution and bearing coal pillar is divided into the three situations: ① when the width of coal pillar W < W_p, only one stress concentration area exists, the bearing capacity of the coal pillar is invalid at this stage, and the lateral mining stress concentration transfers to the roadway solid coal side; ② when the width of the coal pillar W_e ≥ W ≥ W_p, two stress concentration areas appear at this stage, and the coal pillar is in the critical state of self‐bearing stability; ③ when the width of the coal pillar W > W_e, three stress concentration areas are present, and the coal pillar at this stage is in a self‐bearing stable state. Among all these factors, only the size of coal pillar is completely controllable, so the aspects of safe bearing and reserved size design of coal pillar, after estimating the critical size of coal pillar, the coal pillar size design is carried out according to the mine pressure control needs of mining engineering, and the cohesion, internal friction angle, interlayer friction coefficient, and coal seam mining height are improved by artificial technology, so as to realize the resource safe and efficient mining of all kinds of coal seam mining conditions; in the calculation of wide coal pillar size, the advance mining stress concentration at the end of the self‐working face should be taken as the mining load condition, and the reserved size meets the condition of W > W_e, thereby ensuring the stable bearing of the wide coal pillar despite the advanced mining stress concentration during the self‐working face mining; in the calculation of narrow coal pillar size, the lateral mining stress concentration before mining should be taken as the mining load condition and the reserved size meets the condition W < W_p, thereby realizing the effective transfer of mining stress concentration to the roadway solid coal side.

A dynamic coupled elastoplastic damage model for rock-like materials is proposed. The model takes unified strength theory as the strength criterion. To characterize the different damage between compression and tension, two damage variables both in compression and in tension are introduced into the model. The former is represented by the generalized shear plastic strain and volumetric plastic strain and the latter is expressed with the generalized shear plastic strain. Furthermore, the model takes the strain rate effect into account to reflect the strength enhancement under dynamic loading. Because of the difference in plastic hardening between compression and tension, a modified hardening function is adopted. At the same time, the volume strain from HJC model is modified to be consistent with one of continuum mechanics. The developed model is numerically implemented into LS-DYNA with a semi-implicit algorithm through a user-defined material interface (UMAT). The reliability and accuracy of the developed model are verified by the simulation of four basic experiments with different loading conditions. The proposed model was found to be applicable to different mechanical behaviors of rock-like materials under static and dynamic loading conditions.

Protective seam mining is an effective gas pressure relief method in deep mining. Effective theoretical calculation methods in the current studies on the prediction of pressure relief protection effect of interbed coal and rock masses and their distribution laws are lacking. Thus, the evaluation and research with respect to pressure relief effect in protective seam mining relatively lag behind. This situation restricts the engineering feasibility evaluation and decision making in the protective seam mining. Therefore, the influence of upper protective seam mining on the pressure relief protection effect of coal and rock mass between underlying beds was investigated in this study. On the basis of an analysis of concrete engineering projects, a mechanical model was constructed for the pressure relief protection effect of upper protective seam mining on the coal and rock mass between underlying beds. The distribution equation of pressure relief expansion ratio in the underlying protected seam was also derived. The influence laws of main influencing factors on the pressure relief protection effect of the protected seam were revealed as well. In the end, the pressure relief effect was analyzed and verified for the protected seam before and after mining through numerical simulation and similarity simulation test. The pressure relief effect of upper protective seam mining on the coal and rock mass between underlying beds and the distribution characteristics were deeply explored in this study, which could provide a theoretical reference for the decision making in the gas extraction engineering design and pre‐evaluation of extraction effect. Results show that the effective pressure relief zone (expansion rate>0.3%) of the protected seam beneath the goaf is located within the range of approximately 40 m from the coal wall to the rear part. It also presents an approximate “Λ‐shaped distribution characteristic,” that is, it experiences migration and evolution with the advancement in the working face. Moreover, the peak pressure relief lags behind the coal wall on the working face by nearly 10–20 m. In the numerical simulation, the expansion ratio in the goaf also presents an approximate “Λ‐shaped distribution.” Its effective pressure relief zone is the 50 m range from the coal wall to the rear part of the goaf, and the peak value lags behind the coal wall by around 15 m. The theoretical results and numerical simulation results are basically consistent with the physical experiment results. The expansion rates are 1.25%, 1.268%, and 1.32%, respectively. The elastic modulus E of coal seam and interbed spacing H are the main influencing factors of the swelling deformation and are negatively correlated with the expansion ratio. In the actual mining process, E and H of the protected layer can be measured to infer the expansion deformation of the protected layer.