According to the development requirements of green mining of coal resources, it is imperative to improve the extraction rate of coal and the application of safe and efficient mining technology. Pre-splitting and roof cutting technology is widely used in reducing residual coal pillars and safe pressure relief mining, which has become the crucial technology for pillar-free mining methods. Therefore, it is essential to review and discuss the research hotspots, cutting-edge methods, principles of action, and application areas of the development of this technology. Above all, the research data on pre-splitting and roof-cutting development in the past ten years are summarized and outlined. The research’s hot spots are pressure relief technology and gob-side entry retaining technology. Then, the functional forms of pre-splitting and roof cutting technology are discussed and compared, including explosive blasting (directional energy gathering blasting, liquid explosive blasting, and composite blasting), hydraulic fracturing, liquid CO₂ gas fracturing, and mechanized roof cutting (chain arm saw machine and directional cutting roof rig). Through the analysis of field application cases, the application field is divided into three major areas: non-coal pillar mining (gob-side entry driving with narrow coal pillar, gob-side entry retaining with the filling body, completely gob-side entry retaining, and “N00” construction method), pressure relief at working face (thick and hard main roof cracking and end area hard roof cracking), and pressure relief at roadway (gob-side roadway pressure relief and blasting pressure relief technology for roadways). By detailing the process of each application technology one by one, the principle and mode of pre-splitting in each technology are expounded. Finally, the development prospects of pre-splitting and roof cutting in new technical methods, deep pressure relief mining, intelligent unmanned mining, and green and efficient mining are prospected, providing references for similar projects.

In multi-seam mining, the residual coal pillar (RCP) in the upper gob has an important influence on the layout of the roadway in the lower coal seam. At present, few papers have studied the characteristics of the surrounding rock of gob-side entry driving (GED) with different coal pillar widths under the influence of RCP. This research contributes to improving the recovery rate of the extra-thick coal seam under this condition. The main research contents were as follows: (1) The mechanical parameters of the rock and coal mass were obtained using laboratory experiments coupled with Roclab software. These parameters were substituted into the established main roof structure mechanics model to derive the breakage position of the main roof with the influence of RCP, and the rationality of the calculation results was verified by borehole-scoping. (2) Based on numerical simulation, the evolution laws of the lateral abutment stress in the lower working face at different relative distances to the RCP were studied. FLAC3D was used to study the whole space-time evolution law of deviatoric stress and plastic zone of GED during driving and retreating periods with various coal pillar widths under the influence of RCP. (3) The plasticization factor P was introduced to quantify the evolution of the plastic zone in different subdivisions of the roadway surrounding rock, so as to better evaluate the bearing performance of the surrounding rock, which enabled a more effective determination of the reasonable coal pillar width. The field application results showed that it was feasible to set up the gob-side entry with an 8 m coal pillar below the RCP. The targeted support techniques with an 8 m coal pillar could effectively control the surrounding rock deformation.

A large coal pillar (usually more than 90 m) is generally left in place to ensure the stability of main roadway groups, due to its long service lifespan, which commonly also causes a significant loss of coal resources. The design of the width of the protective coal pillar and the control system of the surrounding rock are directly determined by the characteristics of the stress field and the damage mechanism under the influence of the mining activities. However, there are few studies on the effects of the partial-stress boosting and the direction deflection of the stress field on the failure evolution of the surrounding rock (especially in multi-seam mining). In this paper, theoretical analysis and numerical simulation are used to investigate the direction evolution of the maximum principal stress in front of the working face with malposition distances between the upper and lower working faces during the influence of double coal seams mining. Furthermore, a large-scale numerical model is used to study the deviatoric stress evolution of the surrounding rock and the propagation process of the plastic zone in the main roadway group with different widths of protective coal pillars. Then, an asymmetric cooperative anchoring classification method is proposed to strengthen the roadway support, depending on the critical area of the deviatoric stress in the roadway surrounding rock. The peak zone deflection of the deviatoric stress determines the evolution direction of the plastic area, and the peak value of the deviatoric stress presents a typical asymmetric stress boosting on both sides of the roadway. These findings are validated by the on-site ground pressure monitoring results and the practical failure modes of the surrounding rock.

Close-distance coal seams are widely distributed in China, and the mining of overlying coal seams leads to floor damage. To grasp the properties and the fracture spans of the damaged main roof in the underlying coal seam, combining the calculation of the floor damage depth with rock damage theory and the formulas for calculating the first and periodic weighting intervals of the damaged main roof and the instability conditions of the damaged key blocks are obtained. Three interaction stability mechanics models are proposed for key blocks with different properties of the upper and lower main roof, and the instability conditions of the lower damaged key blocks are obtained when the fracture lines overlap. When combined with a specific example, the field monitoring verified the calculation results. The research results are as follows: (1) The first and periodic weighting intervals, horizontal thrust between blocks, and critical load of instability of the damaged main roof are significantly reduced. Still, there are differences in its reduction under different loads, rotation angles, and lumpiness. (2) When the fracture lines of the upper and lower main roofs overlap, the stability of the damaged key blocks is the lowest. There are three linkage stability regions in the critical load curves of the two key blocks. (3) In this case, the damage equivalent of the main roof is 0.397, which belongs to the local damage type. Its first and periodic weighting intervals are 40 m and 16 m, which is 22% and 24% less than when there is no damage. (4) A supporting load of 0.489 MPa is required to maintain the stability of the upper key block, and the lower damaged key block is prone to rotary and sliding instability during the first and periodic weighting, respectively. Thus, the supports need to bear a total of 0.988 MPa and 0.761 MPa to maintain the stability of the two key blocks simultaneously. The ground pressure data monitored on-site is in accord with the calculation results.

Under the condition of fully mechanized top coal caving in close-distance coal seams, the surrounding rock of the stopping space easily loses stability during the withdrawal of mining equipment in the working face because the lower coal seam working face is located under the goaf and the overburden rock has a large range of complex interaction. Field investigation, theoretical analysis, laboratory experiment, similar simulation experiment, numerical simulation, and field industrial tests are used to carry out the research on the stability and control of the surrounding rock in the large section stopping space under the goaf in this paper. The research conclusions are as follows. (1) It is determined that the lower coal seam working face can only stop mining under the goaf, and the reasonable stopping position under the goaf should ensure that the key block fracture line of the main roof is behind the support. (2) The interaction law between the main roof’s key blocks of the upper and lower coal seams is analyzed, and the catastrophic conditions for sliding instability and rotary instability of the main roof’s key blocks of the upper and lower coal seams are obtained. (3) “Anchorage with push and pull equipment-Embedded anchorages and trays” integral anchoring technology is developed. The dimensions of the push and pull equipment are determined. (4) Through numerical simulation of the distribution characteristics of the anchor cable pre-stress field, the asymmetric control scheme of “Partition long and short anchor cables + Integral polyurethane mesh + Embedded anchorages and trays for roof protection” is determined. The rock pressure observation shows that the withdrawal of the working face equipment is implemented safely.