The objective of this work was to investigate the damage characteristics and failure modes of gypsum rock under dynamic impact loading. Split Hopkinson pressure bar (SHPB) tests were performed under different strain rates. The strain rate effects on the dynamic peak strength, dynamic elastic modulus, energy density, and crushing size of gypsum rock were analyzed. A numerical model of the SHPB was established using the finite element software, ANSYS 19.0, and its reliability was verified by comparing it to laboratory test results. The results showed that the dynamic peak strength and energy consumption density of gypsum rock increased exponentially with strain rate, and the crushing size decreased exponentially with the strain rate, both findings exhibited an obvious correlation. The dynamic elastic modulus was larger than the static elastic modulus, but did not show a significant correlation. Gypsum rock fracture can be divided into crack compaction, crack initiation, crack propagation, and breaking stages, and is dominated by splitting failure. With increasing strain rate, the interaction between cracks is noticeable, and the failure mode changes from splitting to crushing failure. These results provide theoretical support for improvements of the refinement process in gypsum mines.

Blasting is one of the most widely used and cost-effective techniques for rock excavation and fragmentation in open-pit mining, particularly for large-scale operations. However, repeated or poorly controlled blasting can generate excessive ground vibrations that threaten slope stability by causing structural damage, fracturing of the rock mass, and potential failure. Evaluating the effects of blast-induced vibrations is essential to ensure safe and sustainable mining operations. This study investigates the impact of blasting-induced vibrations on slope stability at the Saindak Copper-Gold Open-Pit Mine in Pakistan. A comprehensive dataset was compiled, including field-monitored ground vibration measurements—specifically peak particle velocity (PPV) and key blast design parameters such as spacing (S), burden (B), stemming length (SL), maximum charge per delay (MCPD), and distance from the blast point (D). Geomechanical properties of slope-forming rock units were validated through laboratory testing. Slope stability was analyzed using pseudo-static limit equilibrium methods (LEMs) based on the Mohr–Coulomb failure criterion, employing four approaches: Fellenius, Janbu, Bishop, and Spencer. Pearson and Spearman correlation analyses quantified the influence of blasting parameters on slope behavior, and sensitivity analysis determined the cumulative distribution of slope failure and dynamic response under increasing seismic loads. FoS values were calculated for both east and west pit slopes under static and dynamic conditions. Among all methods, Spencer consistently yielded the highest FoS values. Under static conditions, FoS was 1.502 for the east slope and 1.254 for the west. Under dynamic loading, FoS declined to 1.308 and 1.102, reductions of 12.9% and 11.3%, respectively, as calculated using the Spencer method. The east slope exhibited greater stability due to its gentler angle. Correlation analysis revealed that burden had a significant negative impact (r = −0.81) on stability. Sensitivity analysis showed that stability deteriorates notably when PPV exceeds 10.9 mm/s. Although daily blasting did not critically compromise stability, the west slope showed greater vulnerability, underscoring the need for stricter control of blasting energy to mitigate vibration-induced instability and promote long-term operational sustainability.

This research was conducted in order to study the relationship between gas geology (initial gas pressure, initial permeability, and buried depth) and effective extraction radius and to achieve precise borehole layouts. Based on the in-hexagon borehole layout mode, the influence of geological factors on borehole effective extraction radius is quantitatively analyzed. Combined with gas geology, the precise borehole layout mode of gas extraction is constructed. The results show that: Based on the two evaluation indexes of borehole number and area redundancy rate, the optimal implementation scheme of the in-hexagon is selected; that is, when the effective extraction radius is R, the borehole spacing along the coal seam strike is $\sqrt{3}R$, and along the dip is 1.5R. Based on the four evaluation indexes of effective extraction space volume, relative gas emission, cost rate, and gas isobaric surface shape, the relationship between effective extraction radius and initial gas pressure, permeability, and burial depth is matched quantitatively. The effective extraction radius decreases with the initial gas pressure and buried depth and increases with the initial permeability. The effective extraction radius and initial gas pressure have a linear relationship $R=aP+b$, the effective extraction radius and initial permeability have a power function relationship $R=a{k}^b$, and the effective extraction radius and burial depth have a negative exponential relationship $R =a{e}^{-\mathit{bH}}$. The response surface interaction model analysis shows that the buried depth has the strongest influence on the effective radius of gas extraction, followed by the initial gas pressure and the initial permeability. Based on the effective extraction radius as a function of gas geology, the precise borehole layout mode of gas extraction is constructed, which can provide a reference for the construction design of underground gas drilling in coal mines. This will provide a technical guarantee for the efficient mining of gas and promote the sustainable development of gas resources.

To prevent landslides, the slope is a crucial component that needs to be evaluated. Mining activities produce slopes, both natural slopes and artificial slopes, and if a slope is not designed properly, its stability will be adversely affected. The purpose of this study is to determine the stability of a slope supported by sedimentary rocks as the constituent material of the slope. Data processing is carried out using the RS2 Version 11 software and finite element methods (FEMs) by considering the value of the strength reduction factor (SRF) and maximum displacement of the slope. The results obtained for stage 1 show that a maximum displacement of 0 m was obtained, along with a critical value of SRF = 1. A maximum displacement of 0.2 m was obtained in stage 2, with a critical SRF of 1.25. In stage 3, 0.2 m was the highest attained displacement, and the critical SRF was 1.26. A maximum displacement of 0.4 m and a critical SRF of 1.31 were found in stage 4. The maximum displacement in stage 5 was 0.8 m, while the critical SRF was 1.34, and the critical SRF in stage 6 was 1.35, while the maximum displacement was 0.8 m. Finally, the maximum displacement in stage 7 was 1.6 m, while the critical SRF was 1.36. A general pattern emerged from the results of stages 1 through 7. Specifically, the maximum permitted displacement value increased with the critical value of SRF. As the slope moved, it also became more stable, with a big critical SRF. If a slope’s deformation exceeds 1.6 m, it will collapse at a safety factor of 1.36. Furthermore, the contour level showed that the factor of safety (FoS) falls between 1.4 and 4.2. Additionally, as sigma 1 and 3 increase, the resulting FoS value increases as well.