Search Results (71)

To address severe mine pressure disasters induced by the coupling of mining-induced dynamic stress and impact disturbance during close-distance coal seam mining, this paper takes the No. 8 and No. 9 close-distance coal seams in the 119 mining area of a coal mine in Ningxia, China, as the engineering background. Theoretical analysis and FLAC^3D numerical simulation methods were adopted to systematically study the evolution of overburden structure, the manifestation law of mine pressure caused by mining disturbance, and the dynamic response mechanism of roadway surrounding rock under impact load. The findings demonstrate: ① Based on key block theory and elasticity mechanics theory, the stress transfer mechanism of the complete bearing type overburden rock in close-range coal seams was clarified. The calculation model of floor plastic zone depth and additional stress was derived, and the influence mechanism of the bearing state of interlayer rock strata on the stability of underlying coal seam roadways was revealed. ② Comparative numerical simulations of mining schemes revealed that both schemes formed a “goaf pressure relief-workface-coal pillar” load-bearing configuration with “upward subsidence and downward bulging” basin-shaped settlement. Scheme A exhibited significantly increased stress peaks and interlayer plastic zones due to repeated mining-induced stress, substantially elevating the risk of strong mine pressure manifestation and surrounding rock instability. ③ Under 8 MPa cosine impact load with a vibration frequency of 50 Hz (peak particle vibration velocity of 9.57 m/s), compared with the unsupported roadway, the bolt–cable collaborative support system reduced the peak displacement of surrounding rock by over 35% and decreased the shock wave propagation velocity by more than 40%, effectively suppressing the expansion of plastic zones and the transfer of impact energy, while significantly enhancing the impact resistance of the roadway. This study not only provides a systematic theoretical basis for close-distance coal seam mining and rock burst prevention but also offers scientific guidance and technical reference for surrounding rock control and dynamic disaster prevention of roadways in similar close-distance coal seam mining projects, which is of important engineering value for ensuring the safe and efficient mining of underground coal resources. Full article

Rock bursts frequently occur in the fault group area in China, seriously restricting the safe and efficient production of coal mines. Based on field investigation, physical experiments, and numerical simulation, this study investigates the rupture types and spatial evolution of microseismic events during the excavation of working face through fault group areas in the TB Coal Mine, where the hard roof asymmetric is cut by faults. It reveals the cooperative instability mechanism of faults and hard roof, as well as the mechanisms of rock burst. Targeted rock burst prevention measures are proposed, including “roof blasting to cut off dynamic and static load transfer” and “coal blasting to reduce abutment stress”. The results demonstrate the following: (1) during mining in fault group areas, the synchronous activation of faults induces shear-type and high-energy microseismic events and the subsequent movement of hard roof, which has been cut by faults, forms asymmetric parallelograms and symmetric inverted trapezoids, and induces tensile-type and high-energy microseismic events. The synchronous activation of faults and the breaking of the hard roof are identified as the primary reason for high-energy microseismic events. (2) As the fault dip angle approaches 90º, the compressive strength of the fault-segmented hard roof strata decreases. Under synchronous activation of faults, roof failure concentrates in the central, right, and left sections for fault combinations with dip angles of 70° + 70°, 90° + 70°, and 110° + 70°, respectively. (3) Numerical simulations reveal two rock burst mechanisms in faults—hard roof systems: a forward “high dynamic stress and high static stress” type and a rear “low dynamic stress and high static stress “ type, which is consistent with in situ monitoring data. (4) For the three stages in which the 502 working face approaches, passes through, and mines away from the fault group area, a stress relief scheme combining roof blasting and coal blasting is proposed. Compared with the 501 working face, during the mining of the 502 working face, the total microseismic frequency and energy decreased by 71.9% and 87.9%, respectively, and the effectiveness of these measures is verified. Full article

Rock burst is one of the most typical dynamic disasters in the process of coal mining. Energy-absorbing components are key to anti-impact equipment. Exploring the mutual feedback relationship between energy-absorbing components and columns for the prevention and control of roadway rock burst is of great significance. In this study, an arc-shaped energy-absorbing component was designed, and its energy-absorbing characteristics were analyzed. The finite element analysis results of the arc-shaped energy-absorbing component were verified via the crushing test machine. The energy-absorption effect of pre-folded, diameter-expanded, eversion, and arc energy-absorbing columns under impact is compared horizontally. The results show that the supporting force is stable during the crushing deformation of the arc-shaped energy-absorbing component, and the average supporting force measured in the test is 1145.35 kN. Compared with the other three energy-absorbing columns, the arc-shaped energy-absorbing column has a lower emulsion pressure peak and the maximum pressure fluctuation amplitude during the impact process; it also has a better deceleration effect on the quality mass. During the impact process, the influence of the arc-shaped energy-absorbing component on the liquid impact in the column is summarized into three stages, namely approximate elasticity, flexible yield energy-absorbing, and approximate rigidity, which can achieve the peak clipping effect on the liquid impact and improve the impact resistance of the column. Full article

Based on the problem of water and sand inrush caused by the infiltration and failure of the clay layer at the bottom of the loose layer in shallow coal seam mining in eastern China, this study adopts the Particle Flow Code numerical simulation method to conduct multi-physics field coupling analysis. Based on the geological conditions of Taiping Coal Mine in Shandong Province, a two-dimensional water sand clay coupling model was constructed to systematically simulate the entire process of permeability failure of clay layers under different mining crack widths (5–20 mm). The permeability failure mechanism was revealed through porosity distribution, particle contact number, and contact force evolution laws. The numerical simulation results show that with the increase in crack width, the speed of contact reduction is faster, the speed of water and inrush is faster, and the time is shorter. The process of infiltration failure can be divided into two stages: the first stage is the clay infiltration deformation stage, and the second stage is the water inrush and sand collapse stage. In addition, the larger the width of the crack, the greater the contact force, and the shorter the time of infiltration failure and water and sand bursting experienced. The quantitative relationship between the width of mining induced cracks and permeability failure was revealed, and a critical discrimination index for permeability failure in clay layers was established, providing theoretical support for optimizing safe mining parameters and preventing water and sand inrush disasters in porous aquifers. Full article

To alleviate strong strata-pressure bursts during ultra-thick coal extraction, we selected the 26 m number five seam of the Chenjiagou Coal Mine as a full-scale prototype. Three objectives were pursued: (1) elucidate the initiation mechanism of high-energy roof failures under top-coal caving (TCC); (2) quantitatively link the failure sequence of key strata to burst intensity; and (3) deliver field-oriented prevention criteria. A 1:300 physical similarity model and UDEC plane-strain simulations were combined to monitor roof deformation, stress evolution and dynamic response during extraction. Results demonstrate that pressure bursts are driven by abrupt kinematics of the overburden, triggered by sequential breakage of key horizons: the secondary key stratum collapsed at 130 m face advance, followed by the main-key stratum at 360 m. Their combined rupture generated a violent energy release, with roof displacement accelerating markedly after the main horizon failed. We therefore propose two dimensionless indices—the dynamic load factor (DLF) and stress concentration factor (SCF)—to characterize burst severity; peak values reached 1.62 and 2.43, respectively, while pronounced stress accumulation was localized 6–15 m ahead of the face. These metrics furnish a theoretical basis for early warning systems and control strategies aimed at intense rock burst. Full article

During deep mining engineering, coal bodies are subjected to complex geological stresses such as periodic roof pressure and blasting impacts, which may induce mechanical property deterioration and trigger severe rock burst accidents. This study systematically investigated the mechanical characteristics and failure mechanisms of coal under strain rates on two orders of magnitude through quasi-static cyclic loading–unloading experiments and split Hopkinson pressure bar (SHPB) tests, combined with acoustic emission (AE) localization and crack characteristic stress analysis. The research focused on the differential mechanical responses of coal-rock masses under distinct stress environments in deep mining. The results demonstrated that under quasi-static loading, the stress–strain curve exhibited four characteristic stages: compaction (I), linear elasticity (II), nonlinear crack propagation (III), and post-peak softening (IV). The peak strain displayed linear growth with increasing cycle, accompanied by a failure mode characterized by oblique shear failure that induced a transition from gradual to abrupt increases in the AE counts. In contrast, under the dynamic loading conditions, there was a bifurcated post-peak phase consisting of two unloading stages due to elastic rebound effects, with nonlinear growth of the peak strain and an interlaced failure pattern combining lateral tensile cracks and axial compressive fractures. The two loading conditions exhibited similar evolutionary trends in crack damage stress, though a slight reduction in stress occurred during the final dynamic loading phase due to accumulated damage. Notably, the crack closure stress under quasi-static loading followed a decrease–increase pattern with cycle progression, whereas the dynamic loading conditions presented the inverse increase–decrease tendency. These findings provide theoretical foundations for stability control in underground engineering and prevention of dynamic hazards. Full article

After entering deep mining, coal mines often experience various intense dynamic load phenomena due to increasingly complex geological conditions, which can lead to secondary disasters, where it is urgent to identify their sources and analyze their disaster-causing effects. This article takes the 3310 working face in Gu Cheng Coal Mine as the engineering background, and uses theoretical analysis, numerical simulation, on-site monitoring, and other methods to analyze the spatial and temporal distribution of dynamic load events during the mining period of this face. The study classifies dynamic load events based on this background into roof-type, fault-type, and coal pillar-type classes, revealing the differences in the spectra, waveforms, and disaster-causing effects of each class. The results show that the strong dynamic load events are mainly concentrated in the working face roof and fault zone areas. The first principal frequency of the three classes has an estimated boundary between 30 and 60 Hz. The waveform decay coefficients of the roof-type, coal pillar-type, and fault-type strong dynamic load events have average values of 4.53, 1.57, and 1.41, respectively. By adopting the above research methods, a theoretical basis can be provided for the source of dynamic loads, thereby achieving source-based prevention and control of rock burst. Full article

Ensuring the accurate prediction, prevention, and control of coal bursts in isolated working faces is crucial for ensuring safe mining operations. Coal bursts are typically caused by the accumulation of stress and energy released in coal seams and the overlying strata. This study focuses on the 76 isolated working faces at Shanxi Wuyang Mine, employing a combination of theoretical analysis, numerical simulation, and field monitoring. Through theoretical analysis, the study examines the influence of the spatial structure of the overlying strata on support stress and develops corresponding estimation functions. Additionally, bearing strength calculation formulas under varying confining pressures are derived. Numerical simulations are used to validate the effectiveness of borehole stress relief, while field monitoring further confirms the accuracy of the proposed model, leading to the development of the “overall–local” coal burst prediction method. The results demonstrate that the proposed method effectively assesses coal burst risks and, based on different coal burst types, recommends borehole stress relief and roof deep-hole blasting as primary mitigation strategies. These methods were successfully applied to the 76 isolated working faces at Wuyang Mine, yielding conclusions of overall stability with localized instability. This study provides new insights into coal burst prediction theory and offers practical guidance for preventive engineering in isolated working faces, demonstrating substantial engineering applicability. Full article

In view of the high-level, thick, and hard roof in a mine in Shaanxi, it is difficult for existing technology to solve the problem of frequent rock bursts, which are caused by the direct weakening of the whole underground layer. In this paper, a technology for preventing rock bursts using the long-distance drilling and blasting of a thick and hard roof in a high drilling field is proposed. The authors used theoretical analyses, numerical simulations, and other research methods to analyze the mechanisms of pressure relief and load reduction achieved by this technology, determined its layout parameters and layers, and carried out engineering practices in 2412 working faces in a mine in Shaanxi. The results show that the long-distance drilling and blasting technology can achieve the aim of unloading the pressure drop load by arranging a high-level drilling field to achieve the whole-layer presplitting of the thick and hard roof above the working face. According to the orthogonal test method, when using long-distance drilling and blasting under the condition of a high-level roof, the choice of the blasting layer is the biggest factor affecting the change in overburden subsidence. Using the identification basis of the main control disaster causing the layer of overburden, it was determined that 52~67 m above the coal seam of the 2412 working faces was the blasting layer. According to the periodic weighting interval of the working face and the development radius of the fractures in the blasting surrounding rock, the blast hole spacing was determined to be 30 m. After long-distance drilling and blasting, the frequency and energy of micro seismic events were reduced, the entry deformation was reduced compared with the common roof deep-hole blasting technology, and the pressure relief effect of the long-distance drilling and blasting technology was better. These research conclusions can provide theoretical support for the prevention and control of rock bursts during mining production under similar conditions by reducing the load and the unloading pressure on thick and hard roof layers that are difficult to unload from the source. Full article

The stress environments of gob-side roadways (GSRs) are becoming increasingly complex during deep coal mining under thick and hard roofs. This leads to strong strata behaviors, including roadway floor heave, roof subsidence, and even coal bursts. Among them, coal bursts pose the greatest threat to production safety in coal mines. Coal bursts in a GSR strongly correlate with the load characteristics and stress-energy evolution laws of the roadway. This study analyzes the roof structures of double working faces (DWFs) during the initial weighting stage (IWS) and full mining stage (FMS) of gob-side working faces (GSWFs). This study also explores how varying roof structures affect the stability of GSRs. Three-dimensional roof structure models of DWFs and mechanical models of dynamic and static loads superposition on a GSR throughout the IWS and FMS of a GSWF were developed. An analysis identified the primary stress sources affecting the GSR throughout various mining stages of the GSWF. Subsequently, the principle of “three-load” superposition was developed. A novel method was proposed to quantify the stress state in the GSR surrounding rock across different mining stages of the GSWF. The method quantitatively characterizes the load of the GSR surrounding rock. Based on this, the criterion for judging the burst failure of the roadway was established. Numerical simulations are used to analyze the stress-energy evolution laws of the working face, coal pillar, and GSR surrounding rock during the mining process of the GSWF. These findings offer valuable references for studying and preventing coal bursts in GSRs under equivalent geological situations. Full article

Real-time drilling depressurization technology is widely used in the prevention and control of dynamic disasters, such as deep-seated rock burst. However, current coal- and rock-loading tests under drilling conditions seldom account for real-time issues associated with drilling, thus failing to fully reflect the actual stress state of the surrounding rock during the implementation of drilling depressurization technology. Therefore, this study designed and implemented a uniaxial loading scheme for coal samples incorporating real-time-drilling characteristics. The results indicate a significant reduction in the uniaxial compressive strength ($R_{\mathrm{C}}$), elastic energy index ($W_{\mathrm{ET}}$), and impact energy index ($K_{\mathrm{E}}$) of the samples post-drilling. These parameters show a clear decreasing trend with increasing axial stress during real-time drilling. The weakening effect of impact tendency following real-time drilling depressurization is significant, and the depressurization effect is pronounced. The $R_{\mathrm{C}}$, $W_{\mathrm{ET}}$, and $K_{\mathrm{E}}$ of each real-time-drilled sample exhibit a notable decrease with increasing drilling stress, with the reduction rate significantly diminishing after the drilling stress reaches 20% of the peak strength. Full article

The occurrence of rock bursts due to penetrating faults are frequent in China, thereby limiting the safe production of coal mines. Based on the engineering background of a 501 working face in a TB coal mine, this paper investigates stress and energy evolution during the excavation of this working face due to multiple penetrating faults. Utilizing both theoretical analysis and numerical simulations, this study reveals the rock burst mechanism within the triangular coal pillar influenced by the penetrating faults. Based on the evolution of stress within the triangular coal pillar, a stress index has been devised to categorize both the rock burst danger regions and the levels of rock burst risks associated with the triangular coal pillar. Furthermore, targeted stress relief measures are proposed for various energy accumulation areas within the triangular coal pillar. The results demonstrate that: (1) the superimposed tectonic stress resulting from the T6 and T5 penetrating faults exhibits asymmetric distribution and has an influence range of about 90 m in the triangular coal pillar, reaching a peak value of 11.21 MPa at a distance of 13 m from the fault plane; (2) affected by the barrier effect of penetrating faults, the abutment stress of the working face is concentrated in the triangular coal pillar, and the magnitude of the abutment stress is positively and negatively correlated with the fault plane barrier effect and the width of the triangular coal pillar, respectively; (3) the exponential increase in abutment stress and tectonic stress as the width of the triangular coal pillar decreases leads to a high concentration of static stress, which induces pillar burst under the disturbance of dynamic stress from fault activation; (4) the numerical simulation shows that when the working face is 150 m away from the fault, the static stress and accumulated energy in the triangle coal pillar begins to rise, reaching the peak at 50 m away from the fault, which is consistent with the theoretical analysis; (5) the constructed stress index indicates that the triangular coal pillar exhibits moderate rock burst risks when its width is between 73 to 200 m, and exhibits high rock burst risks when the width is within 0 to 73 m. The energy accumulation pattern of the triangular coal pillar reveals that separate stress relief measures should be implemented within the ranges of 50 to 150 m and 0 to 50 m, respectively, in order to enhance the effectiveness of stress relief. Blasting stress relief measures for the roof and coal are proposed, and the effectiveness of these measures is subsequently verified. Full article

Determining the advanced mining influence range of an underground working face is crucial for preventing dynamic disasters, such as coal bursts and gas outbursts. In this study, the occurrence of advanced seismicity before the working face as well as its correlation with the acoustic emission (AE) activity of coal and rocks under axial loading was analyzed. Based on the results, a novel statistical method to determine the advanced mining influence range based on advanced seismicity data was proposed and then validated with a case study. The results show that advanced seismicity is caused by the combined effects of static and dynamic stresses at the working face. This seismicity can be used to assess the mining influence degree of the working face on the advanced coal and rock mass, and determine the advanced mining influence range. Using the novel statistical method, the normalized curves for the total number and total energy of the advanced mining-induced seismicity can be plotted. Then, the advanced mining influence range can be determined using thresholds. The thresholds can be established based on the AE activities observed in coal and rock samples under axial static loading. In the case study in this research, the thresholds for the total seismic number and total seismic energy are 0.076 and 0.052, respectively. The corresponding advanced mining influence ranges are 275 m and 245 m, respectively. Field monitoring confirms an advanced mining influence range of 255 m, which validates the results obtained using the novel statistical method. Full article

Rock burst is a worldwide prevention and control problem, and the main reason for its occurrence is the concentration of stress in the surrounding rock of the coal roadway. Therefore, it is of great significance to realize the rapid and accurate detection of the stress distribution in the surrounding rock of the roadway for the prevention and control of rock burst. Based on the principle of charge induction, this paper adopts a research method combining theoretical analysis and indoor and field tests to carry out a study on the charge induction detection of stress distribution of surrounding rock in coal seam roadways using the self-developed coal rock charge induction monitor. A theoretical analysis of the charge induction intensity in relation to the stress level is carried out. Indoor tests on the law of charge induction for graded loading of large sized coal samples are carried out. Field detection tests of the charge induction law at different drilling depths on the solid coal side and the large coal pillar side of the coal seam roadway are carried out. The results show a positive correlation between the charge signal intensity and the stress magnitude. The induced charge of coal samples has a tendency to increase with the increase in graded loading stress level. The magnitude of the induced charge can reflect the stress level of the coal body. On the solid coal side, the induced charge has a tendency of increasing and then decreasing with the increase in detection depth. The final results are in good agreement with the results of the drill chip method, which better reflects the distribution of the lateral support pressure of the roadway. On the side of the large coal pillar, the induced charge has a tendency to increase, then decrease, and then increase with the increase in probing depth, which is in good agreement with the distribution of lateral support pressure formed in the elastic core area of the large coal pillar. Therefore, the charge induction technology can be used as a fast, non-contact detection means for the partitioning and stress distribution of the roadway enclosure, which can provide guidance for the target prevention and controlling rock burst and for designing roadway support. Full article

To address the frequent occurrence of rock burst disasters in areas with wide coal pillars during mining in the western mining area of China, the wide coal pillar area of the Tingnan coal mine in Shanxi Province was used as the research background. Theoretical analysis, numerical simulation, and field tests were used to establish the mechanical criterion and the energy criterion for the dynamic instability of wide coal pillars. The process and mechanism of wide coal pillar dynamic instability under dynamic and static load disturbances were revealed, and a wide coal pillar rock burst prevention and control scheme was proposed. The results indicated that when the load above a coal pillar reached the stress failure index and the energy failure index was met, the coal pillar reached the critical conditions for rock burst. With increasing static load, the stress, energy, and range of the plastic zone all showed increasing trends on both sides of the coal pillar. Under a given dynamic load, the stress and plastic zone range of the coal pillar significantly increased compared to those without a dynamic load. Under a given static load, the greater the dynamic load, the more likely the coal pillar was to undergo dynamic instability. The evolution of coal pillar dynamic instability was divided into three stages: energy accumulation, local instability, and dynamic instability. When the critical stress and energy conditions for coal pillar dynamic instability are exceeded, rock burst will occur. To reduce the static and dynamic loads of coal pillars, a rock burst prevention and control scheme of energy release and load reduction was proposed and applied onsite. The monitoring results showed that this control plan effectively reduced the stress of the coal pillar and the dynamic load generated by the fracture of the overlying rock layer, indicating safe mining in this area of wide coal pillars. Full article