Rock bursts in roadway groups of deep mining workfaces are more likely to occur due to concentrated static loads, posing significant threats to mining safety and efficiency. In this study, the coal mine roadway groups in western China are taken as an engineering case to investigate the occurrence mechanism of rock bursts in deep mining workfaces caused by concentrated static loads. A novel prevention and control method based on hydraulic fracturing is proposed, and a sensitivity analysis is conducted on key parameters, including in-situ stress, roof firmness coefficient, flow increment, and borehole spacing, to assess their influence on the hydraulic fracturing effect. The results reveal that the failure energy and peak stress of coal pillars within the roadway group gradually increase due to the continuous pressure exerted by the overlying roof, eventually reaching the conditions necessary to trigger rock bursts. The application of a super-long horizontal staged hydraulic fracturing technology transforms the thick, hard roof into a plastic cushion, which absorbs energy and weakens the stress transfer from the overlying roof. This process effectively reduces the stress concentration in the coal pillars. Furthermore, the hydraulic fracturing effect improves with increasing in-situ stress difference and decreasing borehole spacing, while the effect of flow increment is relatively limited. The study also highlights that higher roof firmness coefficients hinder the effectiveness of hydraulic fracturing, as greater water pressure is required for fracture propagation. Field application of the hydraulic fracturing technique, with parameters including an in-situ stress difference (λ), rock strata firmness coefficient, flow rate of 50 m³/day, borehole spacing of 60 m, and a borehole horizontal level of 40–50 m above the roadway, led to a significant reduction in microseismic events and ground audio frequency, demonstrating a remarkable anti-impact effect on-site. This research provides a theoretical framework and practical insights for the prevention and mitigation of concentrated static load-induced rock bursts in similar mining roadway groups.

A novel two-stage mining method, involving large-height mining of the medium layer followed by top-coal caving in the lower layer, has recently been applied to extremely thick coal seams (> 20 m). Throughout the mining process, the top-coal layer undergoes two distinct stages of disturbance and failure, influenced by multiple factors that impact mining efficiency. Based on geological conditions in western China, this study employs numerical simulations to analyze the impact of coal seam burial depth, medium layer thickness, and its location on top-coal movement and pressure distribution under high-intensity repeated mining. The results reveal the formation of a strip-like stress zone and a wide-bottomed inverted funnel displacement field in the top-coal layer during medium layer mining, with maximum stress and subsidence reaching approximately 34.7 MPa and 3 m, respectively. During the lower layer mining stage, crack propagation and coal fragmentation enhance top-coal caving, forming a sawtooth-shaped displacement boundary. Additionally, maximum top-coal subsidence increases to 5.3–7.3 m as burial depth and medium layer mining height increase. These findings suggest that initiating the first mining face in the middle-lower section of the coal seam, where top-coal stress is highest, promotes efficient coal breakage and smooth caving.

The overlying strata of the lower coal seam is easy to be collapsed causing the roof caving accident at the end face of the mining working face under repeated mining in close-distance coal seams. In order to predict the roof instability of the end face, the mechanical model of the granular arch structure is established in this study to further analyze its main influencing factors. The results show that the mining height of the working face, the advancing speed, the distance of coal seams, the tip-to-face distance, the strength of the surrounding rock and the support setting the load of the support are the main influencing factors on the roof caving of the end face. Subsequently, the prediction model of roof instability in the end face under repeated mining is constructed through the radial basis function neural network (RBFNN) and the above main influencing factors are regarded as input layer indexes. Meanwhile, the roof subsidence, coal wall deformation and support load are determined as the output layer indexes. The predicted results are closer to the results of sample tests. Finally, the early warning system, including monitoring and early warning, data query, emergency management, user management, and system settings, is designed to monitor roof conditions of the end face and timely warn the roof accidents. The field application proves that the system has good practical value, which is of great significance to intelligent prediction of coal mine stope disaster and prevent the end face roof disaster under repeated mining and. This will promote the safe and efficient construction of coal mine production.

With the rapid increase in construction projects, the demand for effective landslide prevention has become increasingly urgent

The double wedges sliding along the weak layer of the foundation can be observed on the slope of the waste dump and the sliding body is divided into the active wedge and passive wedge by the weak foundation and the failure surfaces of the waste dump. Because the conventional limit equilibrium slice method cannot reflect the polygonal slip surface of the slope of the waste dump with weak foundation, this study proposed a double wedge calculation method for the slope of the waste dump with weak foundation. The limit equilibrium analysis is performed on double wedges by considering the direction and values of the interaction force between double wedges to obtain the safety factor of the slope of the waste dump. Meanwhile, the quasi-3D double wedges stability analysis method of the waste dump slope with weak foundation is proposed by considering the influence of the geometry and sliding direction of the slope surface on the slope stability. The safety factor of the inverted dump slope is 0.82, the volume of the sliding body is 6.43 million m³, and the main sliding direction is 20° south by east. The shear strain rate cloud diagram of the section is ‘y’ type distribution, and the sliding body is divided into two independent blocks. The safety factor of the sliding body section obtained by the double wedge method is between 0.76 and 0.92, and the closer to the boundary of the sliding body, the greater the safety factor of the section. The quasi-three-dimensional safety factor obtained by theoretical analysis is 0.817. The results show that the calculation results of quasi-3D double wedge are basically consistent with the calculation results of strength reduction method, while the proposed method is simpler. It can be used as a quick method to evaluate slope stability in engineering practice.