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A nonexplosive expansion method is proposed to apply in the field of rock burst prevention and control in coal mine. The principle of the expansive agent reaction was analysed and specimens with one hole and two holes were tested with and without axial stress or an expansive agent. The results showed that the reaction products of the expansive agent exhibited excellent fluidity and did not induce secondary stress concentration within the specimens. The specimens with two holes exhibited more severe and intricate fractures upon introduction of the expansive agent, and the formation of crisscross networks of cracks occurred in both the horizontal and vertical directions. Stress concentration became more pronounced as the distance between the two holes decreased, making the specimen more vulnerable to cracking. A field test was conducted to evaluate the effectiveness of pressure relief through expansion‐induced fracturing in the roadway of 1318 working face. The results from the field test demonstrated that the quantity of coal powder in the deep area significantly decreased after expansion‐induced fracturing and fell well below the early warning index of the drilling cutting method. [ABSTRACT FROM AUTHOR]

The mechanism of rock burst induced by the superposition of dynamic and static loads in multicoal seam mining is unique. To investigate the propagation attenuation law of large‐energy microseismic events and the induced mechanism of rock burst under this condition, this study employs FLAC3D's dynamic module to simulate and analyze the influence of propagation distance, overburden structure in multicoal seam mining, and interlayer plastic zone on vibration wave attenuation. Results indicate that when coal seams are mined at close distances, vibration waves experience significant attenuation while passing through the plastic zone between two layers of coal. At equal attenuation distances, multicoal seam mining structures exhibit greater effects on vibration wave attenuation. Considering differences between rock‐burst induction mechanisms in close‐distance coal seam group mining versus single coal seam mining, a discriminant criterion for rock bursts induced by superimposed dynamic and static loads in multicoal seam mining is established along with a monitoring and early warning method suitable for such conditions. [ABSTRACT FROM AUTHOR]

Deep buried tunnels face high in-situ stress during the construction process, which can easily trigger rock burst. Flexible protective system is an important way to achieve disaster control. Regarding the problem of low toughness and inability to resist instantaneous impacts in the existing flexible protective materials, this study develops a high-strength and high-toughness flexible geogrid based on three deformation modes that may occur when it intercepts explosive blocks. This process considers the selection of polymer fiber filaments, optimization of weaving methods, and improvement of fabric structure. The stress and deformation characteristics of flexible geogrids are analyzed through single-rib tensile, nodal peel, and static jacking tests. The tests showed that mechanical performance of the flexible geogrid can meet requirements for strong and below level rock burst. To analyze the effects of strip strength strip, elongation at break, and nodal strength on the characteristics of jacking deformation, numerical simulations are conducted based on the discrete element method. The results showed that the elongation at break is controlled within 13%, the strip strength is greater than 0.8 GPa, and the nodal strength is increased to 0.075 GPa or above, which can significantly improve the rock-burst protection ability of the flexible geogrid. Finally, the impact energy absorbed by the flexible geogrid reached 74.23 kJ through the falling ball impact experiment, and the engineering application of the flexible geogrid is achieved in the rock-burst section of the Duoji Tunnel on the CZ Railway. [ABSTRACT FROM AUTHOR]

The prediction of rock bursts is of paramount importance in ensuring the safety of coal mine production. In order to enhance the precision of rock burst prediction, this paper utilizes a working face of the Gengcun Coal Mine as a case study. The paper employs a three-year microseismic monitoring data set from the working face and employs a sensitivity analysis to identify three monitoring indicators with a higher correlation with rock bursts: daily total energy, daily maximum energy, and daily frequency. Three subsets are created from the 10-day monitoring data: daily frequency, daily maximum energy, and daily total energy. The impact risk score of the next day is assessed as the sample label by the expert assessment system. Sample input and sample label define the data set. The long short-term memory (LSTM) neural network is employed to extract the features of time series. The Bayesian optimization algorithm is introduced to optimize the model, and the Bayesian optimization–long short-term memory (BO-LSTM) combination model is established. The prediction effect of the BO-LSTM model is compared with that of the gated recurrent unit (GRU) and the convolutional neural network (1DCNN). The results demonstrate that the BO-LSTM combined model has a practical application value because the four evaluation indexes of the model are mean absolute error (MAE), mean absolute percentage error (MAPE), variance accounted for (VAF), and mean squared error (MSE) of 0.026272, 0.226405, 0.870296, and 0.001102, respectively. These values are better than those of the other two single models. The rock explosion prediction model can make use of the research findings as a guide. [ABSTRACT FROM AUTHOR]

Rock-burst prediction is an important issue faced by deep coal mining operations and holds significant significance for ensuring the safety of workers and the sustainable development of mines. This study investigates the crack-propagation behavior of coal and magnetite under uniaxial compression, introduces the characteristic parameter γ within the framework of the flicker noise spectrum (FNS), analyzes the trends in the variation of acoustic emission (AE) b-value and γ-value, and explores the relationship between these two parameters. Experimental results reveal that the proportion of tensile cracks is significantly higher than shear cracks during the loading process, shear failure becomes the main factor as the specimen approaches failure. Both the b-value and γ-value are proven to effectively reflect the damage evolution process of coal and magnetite and exhibit noticeable changes prior to collapse. Specifically, the peak γ-value serves as a precursor indicator for coal and magnetite collapse. Furthermore, this study uncovers a negative linear relationship between the b-value and γ-value. By delving into the feature parameter γ of AE signals, this research offers a new perspective and approach for predicting the instability of coal and magnetite formations and monitoring dynamic hazards. It contributes to a better understanding of the damage evolution process of coal and magnetite formations and provides valuable guidance and reference for preventing and mitigating dynamic hazards associated with coal and magnetite. Highlights: A novel monitoring indicator for coal and magnetite dynamic hazards, known as the γ-value, was introduced within the framework of the Flicker Noise Spectroscopy method. The study revealed a notable negative linear correlation between the b-value and the γ-value derived from acoustic emission signals. Experimental and on-site data have demonstrated that b-value and γ-value can effectively prevent rock burst. [ABSTRACT FROM AUTHOR]

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 C ), elastic energy index ( W ET ), and impact energy index ( K 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 C , W ET , and K 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. [ABSTRACT FROM AUTHOR]

Different coal failure impact types exhibit different damage statuses and failure modes, resulting in distinct signal characteristics of infrared radiation (IR) and acoustic emission (AE). This paper combines IR and AE monitoring methods to innovatively establish coal damage and failure precursor warning models and obtains the IR and AE precursor characteristics for different coal failure impact types. This research shows that there is a good correspondence between IR and AE timing and spatial distribution of different coal impact types. As the impact tendency increases, the intensity of IR and AE signals increases with coal failure, and the AE positioning points and IR high-temperature areas tend to concentrate. The coal body gradually changes from tensile failure to shear failure. The shear cracks in the failure stage of coal with no, weak, and strong impact are 39.9%, 50.9%, and 53.7%, respectively. The IR and AE instability precursor point of coal with no, weak, and strong impact occurred at 55.2%, 66.3%, and 93.4% of coal failure, respectively. After the IR and AE combined instability precursor point, the dissipated energy and combined damage variable increase rapidly, and the coal body will undergo instability and failure. The research results provide a theoretical basis for comprehensive monitoring of coal body failure and rock burst. [ABSTRACT FROM AUTHOR]

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. [ABSTRACT FROM AUTHOR]

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Copyright of Coal Geology & Exploration is the property of Xian Research Institute of China Coal Research Institute and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Geology & Exploration is the property of Xian Research Institute of China Coal Research Institute and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Geology & Exploration is the property of Xian Research Institute of China Coal Research Institute and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Geology & Exploration is the property of Xian Research Institute of China Coal Research Institute and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Geology & Exploration is the property of Xian Research Institute of China Coal Research Institute and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Geology & Exploration is the property of Xian Research Institute of China Coal Research Institute and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

To investigate the influence of varied mechanical parameters of coal mass on rock burst occurrence during deep roadway tunneling, the surrounding coal and rock mass of a deep roadway were taken as the research objects. A geometric model of roadway tunneling was developed using 3DEC numerical simulation software, and the failure characteristics of the coal mass in the roadway side were analyzed based on the rock burst mechanism and stress difference gradient theory for deep mining. The risk of rock burst during roadway tunneling was quantitatively assessed using the change rate of the stress difference gradient (Dgc), thereby elucidating the burst failure patterns of the deep roadway under the influence of varied mechanical parameters. The findings indicate that the coal mass in the roadway side zone is more prone to burst failure due to stress disturbances during deep excavation compared to the coal and rock mass in the roof and floor zones, and that the released kinetic energy and the risk of burst failure are positively correlated with the magnitude of the ground stress. The variation of the mechanical properties of coal mass has a significant effect on the rock burst risk during roadway tunneling. The variation of both internal friction angle and cohesion significantly affects rock burst, with cohesion exerting a greater influence. Conversely, the elastic modulus does not significantly impact the risk. The tendency of bursting in the coal mass is positively correlated with the coefficient of variation (COV) in cohesion and negatively correlated with the COV in internal friction angle. These research findings offer valuable insights for the quantitative assessment of rock burst risk during roadway tunneling. [ABSTRACT FROM AUTHOR]

Rock burst disasters severely restrict the safe and efficient mining of coal. The fundamental cause of their occurrence is the concentration of stress within the coal mass. Stress monitoring in coal seam drilling is widely used as an effective method for rock burst monitoring. However, how to scientifically and reasonably set the critical values of early warning indicators that match the conditions of each mine has always been a key issue restricting the accurate prediction of rock burst by the drilling stress method. This paper adopts a method combining theoretical analysis and field practice to conduct research on the critical values of drilling stress early warning indicators. Based on perturbation response instability theory, a mechanical model for the occurrence of impact ground pressure has been established. Based on the instability theory of disturbance response, a mechanical model for the occurrence of impact ground pressure has been established, leading to the derivation of the expression for the near-field critical stress of impact ground pressure events. The theoretical formula for the critical value of drilling stress early warning indicators was obtained based on the difference between the critical stress of rock burst occurrence and the actual stress in the roadway. This formula includes the mechanical parameters of the coal mass and its propensity for rock burst, roadway support stress, mining depth, stress concentration coefficient, and the initial installation pressure of the stress gauge. They can be determined by the geological and mining technical conditions of each mine. This theoretical formula breaks the uniformity of the critical values for stress warning indicators in various mine drill holes, allowing each mine to scientifically determine its critical value based on its own conditions. This theoretical method has been applied to a high-stress mine in Shanxi, China, and the critical values of drilling stress early warning indicators were obtained. When the monitored stress exceeded the critical value, dynamic phenomena of anchor rod and cable fractures occurred in the roadway roof. The distribution of microseismic events also shifted towards the warning area, and the microseismic monitoring indicators reached the warning values. This confirmed the engineering feasibility of the critical values for drilling stress early warning indicators determined by the theoretical method. [ABSTRACT FROM AUTHOR]

In order to quantitatively calculate the critical depth and critical load of mines affected by rock burst, and to achieve effective prevention and control of rock burst in coal mines, this paper proposes a mechanical model for predicting the occurrence of rock burst in coal mine roadways based on catastrophe theory. Additionally, a theoretical calculation formula for initiating rock burst is derived. The first step was to establish a mechanical analysis model, which directly correlated with the in-situ stress, physical and mechanical characteristics of the coal-rock mass, and engineering structural parameters. Following this, a mechanical instability criterion was derived for the key load-bearing circle within the surrounding rock of the roadway. In the final step, the critical depth and load for rock burst initiation were verified for 25 distinct coal mines in China that were prone to rock burst hazards. The research results demonstrate that the discrepancy between the theoretically calculated critical depth and the actual measured statistical values was less than 35%. In addition, the difference between the theoretically determined critical depth and the value calculated by Pan Yishan was less than 32%. Notably, the ratio of the theoretically calculated critical load to the uniaxial compressive strength of the coal-rock mass ranged from 0.38 to 1.93. This aligns with empirical data on rock burst occurrences, as set out in the engineering classification standards for rock masses. These research outcomes substantiated the practical utility of the proposed theory, thereby laying a robust theoretical groundwork for the quantitative control of rock burst. [ABSTRACT FROM AUTHOR]

Copyright of China Mining Magazine is the property of China Mining Magazine Co., Ltd. and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Abstract The mechanism of rock burst induced by the superposition of dynamic and static loads in multicoal seam mining is unique. To investigate the propagation attenuation law of large‐energy microseismic events and the induced mechanism of rock burst under this condition, this study employs FLAC3D's dynamic module to simulate and analyze the influence of propagation distance, overburden structure in multicoal seam mining, and interlayer plastic zone on vibration wave attenuation. Results indicate that when coal seams are mined at close distances, vibration waves experience significant attenuation while passing through the plastic zone between two layers of coal. At equal attenuation distances, multicoal seam mining structures exhibit greater effects on vibration wave attenuation. Considering differences between rock‐burst induction mechanisms in close‐distance coal seam group mining versus single coal seam mining, a discriminant criterion for rock bursts induced by superimposed dynamic and static loads in multicoal seam mining is established along with a monitoring and early warning method suitable for such conditions.

Abstract A nonexplosive expansion method is proposed to apply in the field of rock burst prevention and control in coal mine. The principle of the expansive agent reaction was analysed and specimens with one hole and two holes were tested with and without axial stress or an expansive agent. The results showed that the reaction products of the expansive agent exhibited excellent fluidity and did not induce secondary stress concentration within the specimens. The specimens with two holes exhibited more severe and intricate fractures upon introduction of the expansive agent, and the formation of crisscross networks of cracks occurred in both the horizontal and vertical directions. Stress concentration became more pronounced as the distance between the two holes decreased, making the specimen more vulnerable to cracking. A field test was conducted to evaluate the effectiveness of pressure relief through expansion‐induced fracturing in the roadway of 1318 working face. The results from the field test demonstrated that the quantity of coal powder in the deep area significantly decreased after expansion‐induced fracturing and fell well below the early warning index of the drilling cutting method.

Drillhole destressing is a local measure in the prevention and control of rock burst. However, existing research has problems such as unclear understanding of the engineering background and drillhole destressing itself, resulting in low credibility of some studies. This article adopts the research method of literature survey and rational reflection, focusing on the definition, the early research history in the prevention and control of rock burst in China, typical engineering background, classification, typical cases of drillhole destressing. Drillhole is an engineering technology that achieves pressure relief through drilling, and destressing drillhole is a borehole constructed for the purpose of pressure relief. Drillhole destressing did not appear in the early research stage of the prevention and control of rock burst in China. After entering the 21st century, research on this issue gradually increased in China. Typical engineering background for drillhole destressing includes roadway excavation, mining, roadway restoration, etc. According to different classification basis, drillhole destressing can be divided into different types. The design of drillhole destressing mainly involves the design of geometric parameters of destressing drillhole, the design of destressing drillhole combinations and the construction design of drillhole destressing. Current research has problems such as incomplete correspondence between drillhole destressing engineering and theory, excessive focus on material properties, unclear distinction about engineering background, and weak research on drillhole destressing on the side of coal pillars. In the future, methods and means that conform to the scientific research paradigm should be selected to conduct targeted, convincing, and refined research.

Abstract The thick and hard roof (THR) is a significant factor that induces rock burst incidents. Traditional deep-hole blasting methods struggle to effectively relieve pressure for the high-level THR on a large scale. Although the horizontal staged hydraulic fracturing technology can crack high-level rock stratas, its impact on roof fracture still needs to be clarified, and there need to be more effective methods to evaluate its pressure relief effect. This paper conducted a comparative engineering test of local hydraulic fracturing pressure relief and load reduction on the working face to address this gap. The key layer theory and mine earthquake distribution were used to identify potential hazards of rock strata and the fracturing strata. Revealed the spatial and temporal evolution rules of microseisms induced by mining work, the failure mechanism of high-energy mine earthquakes, the evolution characteristics of source mechanical parameters, and the evolution characteristics of the stope stress field. The results indicate that after fracturing microseisms during mining presented exhibit a uniform distribution of high frequency and low energy. The fracture fragmentation of the roof rock was reduced, and the range of mining disturbance was minimized. High-energy mine earthquakes induced by mining are transferred to the goaf. Simultaneously, the corner frequency and stress decrease while the rupture radius and the shear component of the failure mechanism increase. The integrity of the THR is compromised, and fracture propagates and slips along the pre-crack. The roof of the working face experiences localized pressure, with decreased periodicity and intensity, reducing the overall static load level of the coal seam under the fracturing area. The research findings serve as a valuable reference for further investigations into the mechanism and risk assessment of hydraulic fracturing in preventing and controlling rock burst.

In view of the current situation that mine impact risk evaluation mainly focuses on the design stage or prior to the implementation of anti-impact measures, and lacks the safety evaluation of mining face after the implementation of anti-impact measures, the types of rock burst are divided into five types: coal type, coal pillar type, roof type, fault type and comprehensive type, taking into account the occurrence location, failure form, load type and focal location; based on the idea of classification prevention and control, a classification weighted evaluation method is proposed, considering the impact of impact risk, roadway surrounding rock, monitoring and early warning system, anti-impact support system, anti-impact rock pressure prevention measures and anti-impact management ability on anti-impact safety; the calculation formula of safety factor of anti-impact is given. The safety level of anti-impact is divided into three levels: safety, basic safety and unsafe. Taking the 31119 working face of a mine as an example, the complete process and result of anti-impact safety evaluation are given. The safety evaluation of some typical mining face with impact risk is carried out by using the classification weighting evaluation method, and the reliability of the evaluation results is verified by combining with the actual impact occurrence.

In response to the challenging problem of deformation and failure of surrounding rock and support structures in roadways under rock burst, a comprehensive research approach involving numerical calculations, field measurements, and laboratory experiments was employed. The study analyzed the characteristics of deformation and failure of surrounding rock and support structures in roadways under strong impact load, investigated the influence of impact load and mining-induced stress on the dynamic mechanical response of surrounding rock and support structures, developed impact-resistant support materials, and proposed a support-unloading coordination control technology, and selected typical rockburst mines for application. The research findings indicate that when the roadway roof is impacted, the maximum particle velocities of the roof, sidewalls, and floor are 1.41 m/s, 0.63 m/s, and 0.25 m/s, respectively. The surrounding rock on the advancing side of the roadway is most affected by the impact, followed by the lateral side, with the retreating side being the least affected. The combined effect of mining-induced stress and impact load results in a larger range of roadway damage, with the plastic zone area reaching 2.1 times the original rock stress. With increasing impact energy, the deformation and peak particle vibration velocities of the roadway increase sharply, and there is noticeable fluctuation in the force on the bolts (cables). Blast destressing has little effect on the anchoring force of the roadway sidewall bolts but significantly affects the anchoring force of the anchor cables, with an average reduction of 26.7% in anchoring force of the anchor cables. The yield strength of high impact toughness anchor rod is 800 MPa, the elongation at break is 20%, and the impact absorption energy is 150 J; The maximum force extension rate of the high elongation anchor cable is 8%, and the tensile strength is 1770 MPa. The coordination control technology was field-tested in a Kuangou coal mine, where the forces on the sidewall bolts ranged from 53 to 84 kN, and the forces on the roof anchor cables ranged from 122 to 219 kN, all within the safe allowable range. The shallow and deep separation layers of the roof have not occurred, and even in the event of a large energy microseismic event, the support system remained stable.

The observability of underground engineering, such as coal mining, is extremely limited, which makes it difficult to obtain reliable pre-analysis basis for many complex engineering problems. Especially in face of new engineering environment brought about by deep conditions and high-intensity mining, the applicability of the existing theoretical system and technical solutions is facing severe challenges. Based on the technical characteristics of the digital era, a digital rock mechanics solution that integrates data laws, mechanical logic and engineering experience is proposed. The solution makes full use of the objective characteristics of real engineering laws contained in engineering data, and gives a specific mode of underlying integration among theoretical achievements, engineering experience and data law mining. Through the non-damaged inference of mechanical parameters of coal specimens, the feasibility of the above technical path is verified, and the data mining and utilization with both reliability and interpretability are realized. Furthermore, a digital rock mechanics theoretical system with rock physical properties, geological structure and spatial stress as the basic elements is established, and a general operation mode of digital rock mechanics underground engineering with the steps of “engineering sketch, physical knowledge embedding, professional knowledge integration, decision generation” is proposed. The reliability of the digital rock mechanics solution is verified by the application in complex engineering, which provides a practical implementation path for the digital transformation and intelligent upgrading of underground engineering such as coal mining.

To reveal the energy accumulation law of surrounding rock in coal thickness change area and the influence of mining direction on it, this study takes the 17# coal seam of Hegang Junde Coal Mine as the research background, and takes the combined rock strata as the object to discuss the stress transfer mechanism of coal thickness change area. The relationship between the original rock stress and the thickness ratio and elastic modulus ratio of coal rock in the coal thickness change area is revealed, and the influence of mining direction on the mining stress and energy accumulation law in the coal thickness change area is clarified. The following conclusions are obtained: due to the difference in strength (stiffness) between rock and coal, the principal stress distribution in the coal thickness change area is uneven. The peak stress appears in the coal thickness thinning area, and the valley stress appears in the thickening area. The larger the range of coal thickness change and the smaller the elastic modulus ratio of coal and rock, the greater the peak stress and the smaller the valley stress. Under the influence of mining stress, the advance abutment pressure and the abnormal stress of coal thickness change are superimposed in the area of coal thickness change, and the stress concentration of surrounding rock is high. From thin to thick mining, there is an impact danger zone due to the superposition of the advance abutment pressure zone and the stress anomaly zone in front of the working face, and there are two impact danger zones from thick to thin mining, so it is more reasonable to mine from thin to thick. In view of the mining stress and energy accumulation law in the coal thickness change area, the prevention and control ideas of regulating the surrounding rock structure, surrounding rock stress, and increasing impact resistance energy are put forward, the engineering application is carried out, and the application effect is good.

In view of the frequent occurrence of rockburst in the coal pillars of high-stress roadways in deep underground coal mines in China, the types and mechanisms of rockburst in coal pillars are studied by means of field investigation, theoretical analysis, numerical simulation and engineering test. A new method of preventing rockburst by releasing energy and modifying the coal pillars of high-stress roadways is put forward. Based on the method proposed, a series of research work such as the principle research, technical system design and underground industrial test are carried out. Based on the main force source types of coal pillar rockburst in roadways, combined with many typical cases of rockburst, the coal pillar rockburst in roadways in China is divided into three types: basic stress superposition type, additional stress induced type and self-creep instability type. The mechanical models and engineering criteria of the occurrence mechanism of coal pillar rockburst in roadways are established and deduced respectively, which reveals its occurrence mechanism. In this paper, a method of preventing rockburst by combining fracturing with grouting in high-stress roadway in deep mine is proposed. The principle of preventing rockburst is embodied in two aspects: “energy release” and “modification”. The underground industrial test shows that the crack initiation pressure of elastic core coal is about 40 MPa, and the overall maximum fracturing width is 19.5 m. The field monitoring proves that the method proposed can reduce the energy accumulation level of roadway coal pillar and increase the overall impact threshold of roadway coal pillar. A complete technical system of coal pillar energy release modification and erosion prevention in high stress roadways of deep underground coal mines is designed, including the “evaluation-inspection-check” method of impact risk of isolated coal body, the energy release technology of elastic nuclear hydraulic fracturing and the whole grouting reinforcement modification technology. The “vibration-stress-displacement” three-field monitoring equipment for the elastic nuclear fracturing of coal pillars in roadways and the PS-Z-A/B type inorganic two-component quick-setting integral grouting modified material have been developed.

Rock burst disaster has become a major safety problem in underground coal mining worldwide. Using the true triaxial electro-hydraulic servo loading test system, the stress-strain law, failure characteristics and acoustic emission activity law of coal-rock combination, deep coal body and shallow coal body with three different stress types are studied. The strength criterion relationship of deep and shallow coal bodies is established, the energy release characteristics of deep coal body are analyzed, and the instability mechanism of deep coal body is revealed. The results show that the coal samples with different stress types are affected by the stress environment and structural differences. The shallow coal samples of the roadway rib and the coal-rock combination samples of the roadway rib under high σ2 confining pressure all produce rapid stress drop after the peak stress, and the brittle failure and impact are strong. In the failure mechanism, the shallow coal sample of the roadway is mainly controlled by the tensile-shear coupling mechanism. The deep coal sample of the roadway is mainly controlled by the shear mechanism. The area near the constraint surface and the free surface of the coal-rock combination sample is controlled by the shear and tension-shear coupling mechanism respectively. Under the condition of same confining pressure, the shallow coal body of the roadway is more likely to be destroyed. Based on the source of the impact energy release of the roadway coal rib, the instability model of the coal body in the deep roadway is established, and the instability mechanism of the coal body in the deep roadway is revealed.

Abstract Rock burst disasters severely restrict the safe and efficient mining of coal. The fundamental cause of their occurrence is the concentration of stress within the coal mass. Stress monitoring in coal seam drilling is widely used as an effective method for rock burst monitoring. However, how to scientifically and reasonably set the critical values of early warning indicators that match the conditions of each mine has always been a key issue restricting the accurate prediction of rock burst by the drilling stress method. This paper adopts a method combining theoretical analysis and field practice to conduct research on the critical values of drilling stress early warning indicators. Based on perturbation response instability theory, a mechanical model for the occurrence of impact ground pressure has been established. Based on the instability theory of disturbance response, a mechanical model for the occurrence of impact ground pressure has been established, leading to the derivation of the expression for the near-field critical stress of impact ground pressure events. The theoretical formula for the critical value of drilling stress early warning indicators was obtained based on the difference between the critical stress of rock burst occurrence and the actual stress in the roadway. This formula includes the mechanical parameters of the coal mass and its propensity for rock burst, roadway support stress, mining depth, stress concentration coefficient, and the initial installation pressure of the stress gauge. They can be determined by the geological and mining technical conditions of each mine. This theoretical formula breaks the uniformity of the critical values for stress warning indicators in various mine drill holes, allowing each mine to scientifically determine its critical value based on its own conditions. This theoretical method has been applied to a high-stress mine in Shanxi, China, and the critical values of drilling stress early warning indicators were obtained. When the monitored stress exceeded the critical value, dynamic phenomena of anchor rod and cable fractures occurred in the roadway roof. The distribution of microseismic events also shifted towards the warning area, and the microseismic monitoring indicators reached the warning values. This confirmed the engineering feasibility of the critical values for drilling stress early warning indicators determined by the theoretical method.

Coal mining engineering involves complex stress structure time effect issues, which are closely related to the occurrence and prevention of rock burst disasters. Therefore, this study conducts experimental research on the stress structure time effect of impact coal. Coal rock specimens were obtained from three coal mines in Shaanxi, Gansu, and Inner Mongolia. The stress-time effect of loading rate was studied through coal uniaxial compression static-dynamic load and creep tests. The stress-time effect under the condition of coal in situ release was investigated using a long-term segmented test method. Water immersion tests on coal and structural time-effect studies on coal-rock combinations were conducted. The experimental results indicate that there is a positive correlation between the uniaxial compressive strength of coal and the loading rate, and a negative correlation with the duration of in-situ relief and immersion time. The combined mode of coal-rock structure time effect significantly affects its energy and strength properties, while impact coal exhibits both stress and structural time effects. According to the experimental results, a plastic limit structure is proposed to explain the irreversible deformation of coal-rock mass under stress. Furthermore, by constructing a time-dependent model for plastic strain in coal under impact conditions and comparing the numerical approximations of non-elastic strain and irreversible strain, the plastic limit strain characteristics of shock-induced coal were verified.

Aiming at the influence of hard and thick roof fracture and structural evolution on the impact of working face, taking the hard and thick roof condition of 63 upper 05 working face in Dongtan Coal Mine as the engineering background, the critical condition and structural evolution process of hard and thick roof fracture and instability are studied by theoretical analysis and numerical simulation, and the dangerous area of working face impact is divided. It is concluded that the thickness and strength of the roof are proportional to the vertical load at the end of the block breaking. The structure of immediate roof and main roof gradually evolved from “V” type to “U” type. The first and periodic steps of the double hard thick roof increase obviously compared with the conventional condition, and the evolution of the basic roof structure is more retarded than that of the direct roof. The vertical stress of the roof evolves in the form of stress arch, and many stress concentration phenomena are formed in the mined-out area during mining. Based on the above research results, the impact danger area of 63 upper 05 working face was divided and the accuracy was verified by microseismic monitoring technology.

We conducted a study to evaluate the potential and robustness of gradient boosting algorithms in rock burst assessment, established a variational autoencoder (VAE) to address the imbalance rock burst dataset, and proposed a multilevel explainable artificial intelligence (XAI) tailored for tree-based ensemble learning. We collected 537 data from real-world rock burst records and selected four critical features contributing to rock burst occurrences. Initially, we employed data visualization to gain insight into the data's structure and performed correlation analysis to explore the data distribution and feature relationships. Then, we set up a VAE model to generate samples for the minority class due to the imbalanced class distribution. In conjunction with the VAE, we compared and evaluated six state-of-the-art ensemble models, including gradient boosting algorithms and the classical logistic regression model, for rock burst prediction. The results indicated that gradient boosting algorithms outperformed the classical single models, and the VAE-classifier outperformed the original classifier, with the VAE-NGBoost model yielding the most favorable results. Compared to other resampling methods combined with NGBoost for imbalanced datasets, such as synthetic minority oversampling technique (SMOTE), SMOTE-edited nearest neighbours (SMOTE-ENN), and SMOTE-tomek links (SMOTE-Tomek), the VAE-NGBoost model yielded the best performance. Finally, we developed a multilevel XAI model using feature sensitivity analysis, Tree Shapley Additive exPlanations (Tree SHAP), and Anchor to provide an in-depth exploration of the decision-making mechanics of VAE-NGBoost, further enhancing the accountability of tree-based ensemble models in predicting rock burst occurrences.

The coupling effect of static and dynamic stress in fault structure areas exacerbates the risk of rock impact in underground fault areas. The stress distribution law and impact dynamic load response features of roadway surrounding rock in fault structure areas have significant peculiarities. At present, research on the impact of fault structures mainly focuses on the vicinity of the working face, but there is little research on the impact damage of roadways in fault structures. Taking the roadway in a deep buried high stress fault structure area of a mine in Shaanxi Province as the engineering background, the mechanical features of deformation and failure of the roadway surrounding rock in the fault structure area are analyzed. ① There is a significant stress barrier effect on the fault plane, and there are two special stress zones near the normal fault, namely the stress concentration zone in the hanging wall and the stress reduction zone in the lower wall. Due to the influence of the fault plane, the static load concentration stress of the roadway side shows an asymmetric distribution feature. The stress concentration on the side far from the fault plane is greater than that on the side near the fault plane, and the risk of impact damage to the surrounding rock of the roadway on this side increases. ② The fault plane has a significant barrier effect on the transmission of stress waves, and the dynamic load response of the hanging wall of the normal fault is greater than that of the lower wall. Due to the asymmetric distribution of stress on the two sides of the roadway, the dynamic load response of the right side is significantly greater than that of the left side. Based on the above features, a collaborative anti impact control technology of 'unloading (large diameter drilling pressure relief) - support (stepped reinforcement into layered energy absorption and anti impact support)' is proposed for the surrounding rock of the fault structure area roadway. The engineering test results show the following points. ① After adopting the 'unloading support' collaborative anti impact treatment measures for the roadway surrounding rock , the stress concentration areas of the two sides of the roadway are transferred to the deep part of the surrounding rock by 3-5 meters. The stress peak value is reduced by 18.5%-20.3%, and the stress concentration degree of the roadway surrounding rock is significantly reduced. ② Before the implementation of the 'unloading support' collaborative anti impact treatment measures, the deformation of the roadway roof, floor, and two sides are 856, 334, 325, and 567 mm, respectively. The deformation and damage of the roadway surrounding rock are severe. After adopting the 'unloading support' collaborative treatment measures, the deformation of the roadway surrounding rock decreases by 35.69%-62.03%, and the stability of the roadway surrounding rock is enhanced. ③ The coal powder content in the borehole is also significantly lower than the critical coal powder content, and the dynamic power of the roadway surrounding rock is reduced.

Based on the rock burst occurrence of the south water sump connecting roadway in the seventh mining area of Zhaolou Coal Mine, the mechanism of creep instability rock burst of high stress isolated coal immersed in water was studied by means of theoretical analysis, numerical simulation and field measurement. The main conclusions are as follows: the bearing stress of isolated coal immersed in water mainly presents the creep evolution law from “bimodal-shaped” to “saddle-shaped”, then to “platform-shaped” and finally to “unimodal-shaped”. Based on the burgers model, a water damage variable was introduced to establish a mechanical model for estimating the concentrated stress in the elastic zone of isolated coal, revealing the creep instability impact mechanism of high stress isolated coal immersed in water. Under the continuous weakening effect of water accumulation in the mine sump, the strength of isolated coal gradually decreases until residual strength, and the bearing capacity also decreases, leading to the transfer of stress concentration from the plastic zone to the elastic zone. When the concentrated stress in the elastic zone exceeds its instability threshold, it is easy to induce rock burst. Based on this, the prevention and control measures for the creep instability impact of high stress isolated coal immersed in water were proposed: reducing the average bearing stress of the isolated coal, increasing the comprehensive compressive strength of the isolated coal, and filling the tunnels near the isolated coal when necessary.

Abstract The study of rock burst tendency of rock masses with different sizes plays a key role in the prevention of rock burst. Through theoretical analysis, it is proposed that uniaxial compressive strength and deformation modulus ratio are the key mechanical parameters affecting rock burst occurrence. In order to find out the size effect of uniaxial compressive strength and deformation modulus ratio, theoretical analysis and uniaxial compression experiment are carried out on rock samples with different heights, different cross-sectional areas and different volumes. The results show that the smaller the uniaxial compressive strength is, the larger the deformation modulus ratio is, and the more likely rock burst are to occur. On the contrary, rock burst is still not easy to generate. The uniaxial compressive strength of rock samples with different heights, different cross-sectional areas and different volumes increases with the increase of rock sample size. The deformation modulus ratio of rock samples with different heights and different volumes shows an upward trend on the whole, while that of rock samples with different cross-sectional areas shows a downward trend on the whole. The fracture forms of rock are analyzed using the energy conversion law in the process of deformation and failure for three kinds of rock with different shapes and sizes.

In order to prevent damage to the support equipment caused by rock burst, a torsion plate energy absorption member is designed to be used with hydraulic columns. Explicit dynamic simulation is used and compared with the tests. The equations of structural parameters to the mean value of the reaction force (Fmean), specific energy absorption (SEA), and mean squared deviation (σ) are constructed based on the response surface method. The structural parameters are optimized using MOPSO. The results indicate that the torsion plate exhibits excellent energy absorption characteristics. The error rate of the simulation compared to the test is less than 5%, indicating a high level of simulation accuracy. All three sets of response fitted equations have R2 above 0.97, indicating a good fit. The three evaluation indexes demonstrate a trend of increasing then decreasing as the torsion angle increases. The wall thickness is positively correlated with the influence of each parameter index, while the effect of fillet radius is reversed. The maximum relative error of the predicted value is 9.50%. After optimization, the Fmean is effectively increased by more than 78%, the SEA is increased by more than 24.72%. The σ is increased by 8.31%, the optimization effect is remarkable.

Taking the 11–3106 working face of Menkeqing Coal Mine as the research background, a fracture model of the structure covered by thick and hard overburden at the 11–3106 working face was established. The mechanism of rock burst in the open roadway of narrow coal pillar with multiple key layers under the deep thick and hard roof mining condition with high strength was studied. The three-level collaborative pressure relief prevention technology of “drilling, blasting and pressuring” was proposed and applied in the field. Seismic CT, microseismic monitoring, support pressure monitoring and surface subsidence observation were used to verify the effect of pressure relief. The results show that there are high and low key layers overlying the working surface of 11–3106. When the two key layers break, the overlying rock may move in a wide range, and the surrounding rock of the open roadway is greatly affected by the dynamic load disturbance. The low bearing capacity of the narrow coal pillar makes the overburden load transfer to the solid coal side of the roadway and the stress increase on the solid coal side, which is the main area of impact initiation. Under the superposition of the lateral fixed abutment pressure and the mobile abutment pressure of the working face, the stress concentration degree of the solid coal side increases continuously. Affected by the periodic breaking of the hard roof, the “dynamic and static” loading is easy to induce the impact of the open roadway. Hard roof is not only the high static load source in the 11–3106 working face system, but also the source of strong dynamic load outside the system. For the prevention of rock burst at the 11–3106 working face, the measures should be taken from three aspects at the same time: ① the coal body is weakened to form a buffer and isolation zone for impact; ② the low pre-cracking roof is weakened to reduce the density of coal rock mass and destroy the energy storage conditions inside and outside the system; and ③ the high pre-cracking roof is artificially induced to break the key layer in the middle and high levels. The three-level collaborative pressure relief prevention and control technology of “drilling, blasting and pressuring” was developed. After the on-site implementation, the impact dynamic load caused by the breaking of high and low thick hard roof was reduced, the stress concentration in the high stress zone in the coal body was weakened, the high energy microseismic events were significantly reduced, the surrounding rock activity during the mining of working face was reduced, the compressive strength of working face roof was reduced, and the ground settlement value was increased by 200%. The results show that the three-stage coordinated pressure relief of “drilling, blasting and pressuring” has a good effect of “energy release and impact reduction”, which can provide a reference for the rock burst prevention and control of narrow coal pillar working face under similar thick hard roof conditions.

In order to understand the factors leading to rock burst in steeply-inclined coal seams, based on the mining layout and on-site microseismic monitoring data, the distribution of mine microseismic events and previous rock burst events are deeply analyzed from time and space, and the evolution law of disaster factors under different mining depths is studied. The study shows that the rock burst in steeply-inclined coal seams is mainly caused by stress concentration on the working face due to the hanging roof and intermediate rock pillars, as well as the dynamic disturbance induced by the fracture of the coal and rock mass. The daily advance rate, mining depth, and remaining coal pillars are strongly correlated with the occurrence of rock burst events, and are the main factors leading to rock burst in the mine. During shallow mining, the “prying force” effect of the intermediate rock pillars caused by the horizontal tectonic stress and the self-weight stress on the coal-rock mass is the main factor causing instability and failure of the coal-rock mass. After entering deep mining, the “extrusion” effect of the hanging roof is superimposed on the “prying force” effect of the intermediate rock pillars, and the factors leading to rock burst change to the combined “extrusion + prying force” effect of the roof and pillars.

Protective seam mining is widely used as an effective method for preventing and controlling rock burst, the protective layer itself also has the risk of strong mine earthquake and rock burst. Taking the protective seam of the oil shale working face in Haishiwan Coal Mine, which encountered with the strong mining earthquake during the mining process, as the engineering background, the focal mechanism and main control factors of the strong mining earthquake were analyzed through numerical simulation and theoretical analysis. The research results show that the strong mining earthquake in the oil shale working face are mainly caused by compression shear failure under stress concentration. The mining of the protected coal seam working face has a more significant impact on the stress and energy distribution in the floor strata of the protected seam. There are significant influences on the occurrence of strong mine earthquake events, such as hard roof strata, thin rock strata on floor and thickness variation between protective layer and protected layer.

Irregular upper gobs may increase the risk of rock burst incidents in coal mines. Investigating the rock burst mechanism in the complex mining spatial structure of multi-seam rock bursts is essential for ensuring the safe extraction of coal resources. This study investigated the distribution characteristics of static and dynamic loads on the W1123 working face through a comprehensive approach utilizing theoretical analysis, numerical simulation, and on-site monitoring. The mechanism of rock bursts was analyzed and targeted preventive measures were proposed and applied on-site. The results indicated that the coupled effects of the overlying knife-shape-like gob and thick hard roofs exhibited regional characteristics in the distribution of static and dynamic loads beneath the W1123 working face. When the working face was solely influenced by the “handle” effect of the knife-shape-like gob or concurrently affected by both the “handle” and “blade body”, the lower part of the working face and the coal pillar experienced concentrated static loads and intense dynamic load disturbances. The surrounding rock in the upper gob also demonstrated concentrated static and dynamic loads, with the proximity to the boundary of the knife-shape-like gob resulting in higher dynamic and static loads. When only affected by the “blade body”, the upper surrounding rock of the working face experienced concentrated static loads and was disturbed by dynamic loads. The area of concentrated static loads and dynamic load disturbances in the working face coincided, leading to the manifestation of rock burst from the superposition of dynamic and static loads. Based on the structural characteristics of the knife-shape-like gob, the rock burst types during the mining period of the W1123 working face were sequentially classified as strong dynamic load type, high static load, strong dynamic load type, and high static load type. The risk of rock bursts assessed through on-site monitoring and numerical simulation matched the theoretical analysis outcomes. Rock burst prevention measures were optimized based on the distribution characteristics of static and dynamic loads on the W1123 working face. To mitigate the stress and energy concentration on the roof and coal body, we have enhanced the treatment height and crushing degree of the thick, hard roof while also intensifying coal body blasting and pillar cutting. Following the implementation of these optimized measures, the average event energy released by the surrounding rock of the working face decreased by 54%, and significant energy events were substantially reduced, leading to a significant reduction in the rock burst risk. This study can provide valuable insights for rock burst prevention in mines with irregular upper gobs.

The variation of coal seam thickness is one of the important factors affecting the pressure relief of large-diameter drilling hole. We have established an elastic model of pressure relief of large-diameter drilling hole based on the principle of equivalent elastic modulus. Taking two working faces of a mine as engineering background, we have studied key issues of large-diameter drilling hole regarding coal seam thickness, including the stress distribution characteristics and energy release laws, and the pressure relief and energy release mechanism. Several phenomena of engineering significance have been revealed by our research. Firstly, when the parameters of large-diameter drilling hole are the same, the stress of primary rock, tangential stress and plastic zone range of thin coal seam are greater than those of thick coal seam. Secondly, the peak stress around the large-diameter drilling hole decreases nonlinearly with the increase of coal seam thickness. When the coal seam thickness increases from 1 m to 9 m, the peak stress drops by 12.9%, from 23.2 MPa to 20.2 MPa. Likewise, the pressure relief and energy release rate also decreases with the increase of coal seam thickness. When the coal seam thickness increases from 1 m to 9 m, the pressure relief and energy release rate of large-diameter drilling hole decreases from 68.7% to 45.8%, indicating that better pressure relief effect can be achieved in thin coal seam. Moreover, when the parameters of large-diameter drilling hole are the same, the increase of coal seam thickness causes decrease of the tangential stress and thereby results in the decrease of the range of coal fragmentation zone and plastic zone, and the amount of the released elastic energy, which reduces the pressure relief energy release rate. The coal seam thicknesses of working faces 1208 and 1203 of a certain mine are 9.08 m and 4.95 m, respectively. Although the same set of large-diameter drilling hole parameters is adopted for these two working faces, the one with thicker coal seam suffers from more serious surrounding rock deformation and failure. By optimizing the parameters of large-diameter drilling hole based on our research, the average amount of drilling cuttings decreased from 2.48 kg/m to 1.76 kg/m. Therefore, we suggest that as the coal seam thickness increases, measures such as reducing drilling spacing and increasing drilling diameter should be considered in order to guarantee the expected pressure relief effect.

In this paper, true triaxial loading and unloading tests under different principal stress conditions were conducted to study the mechanical properties, failure characteristics, and energy evolution laws of coal bodies that are prone to ejection under different loading and unloading paths, aiming to reveal the ejection mechanism of the surrounding rock of roadway that is prone to ejection after excavation and unloading. The results are as follows. Firstly, the damage of coal samples that are prone to ejection was more severe under high stress single-sided unloading than under loading; when the axial pressure is 90% of the peak value, a large shear crack appeared on the surface of the coal samples, the axial strain rate is high, its own damage degree is large, and the ejection phenomenon was obvious; when the axial pressure was 80% and 70% of the peak value, respectively, the coal samples did not undergo overall macroscopic damage, the axial strain rate is low, and the degree of damage is small, only generated tensile cracks on the free face, and the ejection phenomenon was not obvious. Secondly, the increase of the second principal stress had a reinforcing effect on coal samples that are prone to ejection within a certain range, the internal composite cracks of coal samples that are prone to ejection first increase and then decreased, and the failure mode changed from shear failure to tension-shear composite failure, and finally developed into splitting failure; the intensity of ejection first increased and then decreased. Thirdly, in the process of high stress unloading failure, elastic energy was converted into dissipated energy and released instantly, and the proportion of elastic energy increased sharply, causing the fragments being ejected from the mother body with energy; there were obvious cross mesh cracks in the free face, and the ejection phenomenon was obvious. Fourthly, for coal samples that are prone to ejection under high stress after unloading, the tension-shear cracks rapidly expanded and connected, resulting in tension-shear composite failure, and the ejection phenomenon was obvious; Fifth, tension-shear cracks appear in ejectable coal samples of after unloading, RA increased sharply and AF decreased continuously, AE showing the characteristics of high energy, high meter, high amplitude and high frequency. The study results can provide reference for the on trol of micro impact phenomena in unloading surrounding rock during excavation of similar underground engineering.

In response to the challenging task of accurately identifying the main hazard-causing layer of overlying strata in the coal mine surface hydraulic fracturing construction, this study focuses on the industrial test of ground hydraulic fracturing at the 401102 working face of the Mengcun Coal Mine. The research is conducted using the methods of theoretical analysis, microseismic monitoring, and on-site investigation to reveal the dynamic disaster mechanism of mine earthquakes and rock bursts induced by the movement of thick and hard overlying strata in the coal mines. The relationship between the movement characteristics of thick and hard overlying strata based on a three-zone structure loading model of overlying strata and induced dynamic disasters is analyzed, and a prediction model for mining seismic energy and an estimation model for equivalent additional stress in mining areas based on the movement state of key layers are established. A coal mine identification technology for the main hazard-causing layer of overlying strata is proposed based on the K-means clustering algorithm and the elbow rule. The construction layer for hydraulic fracturing is determined, and an industrial test is carried out on-site. The effectiveness is verified based on the microseismic monitoring data and theoretical analysis results, leading to the following conclusions. In the Mengcun Coal Mine’s 401102 working face, both the key stratum responsible for rock bursts and mine seismic activities can be traced to the R9 key stratum of the Anding Group, situated 66 meters away from the coal seam. The primary fracturing movement of this critical stratum R9 imparts an equivalent supplementary disturbance stress value of 7.23 MPa, with the seismic energy liberated by this initial rupture motion quantifying to 6.08×105 J, thereby indicating a pronounced susceptibility towards catastrophic occurrences. After fracturing the key layer which induces mining earthquakes and rock bursts, the theoretical value of the mine earthquake energy is reduced by 94%, and the theoretical value of the equivalent disturbance stress of the working face is reduced by 76%. High-energy microseismic events above the working face with an energy of 5×103 J show a noticeable upward trend, with an upward movement of approximately 15 m. The frequency ratio of microseismic events with an energy level of 103 J or higher significantly decreases from 60.39% to 17.89%, and the maximum microseismic event energy decreases from 6.65×105 J to 9.75×103 J. The proportion of microseismic events with an energy level of 102 J and below significantly increases from 39.61% to 82.11%.

In order to study the mechanics, acoustic emission (AE) and electromagnetic emission (EME) response law of bursting liability coal at different loading rates, uniaxial compression tests were carried out on coal mass from Konggu Coal Mine. The corresponding relations among mechanical properties, AE and EME signals in the process of coal failure under loading were analyzed, and the energy evolution law of coal failure with bursting liability under loading rate was discussed. The results show that within a certain range of loading rate, the higher the loading rate, the higher the compressive strength and peak load of bursting liability coal, and the shorter the time for coal to reach the peak load. Under different loading rates, the mechanics, AE and EME signals of coal samples can be well corresponded. When the loading rate is low, the number of blocks destroyed of coal sample is large and the block size is relatively small, and the blocks are mainly scattered around the test platform. When the loading rate is high, the number of damaged blocks is relatively small and the block size is relatively large, and the blocks are far away from the test bench. When loading at a low rate, the internal cracks in coal can be fully developed and connected, and the energy release rate is relatively uniform in the process of loading and failure of coal sample. In the case of high loading rate, the energy release rate of coal sample in the loading process is much smaller than that in the moment of failure. Combining the above test results with the actual situation of the working face, it can be concluded that the total energy stored in the coal of fast mining increases and the threshold of impact decreases compared with that of slow mining. Therefore, under the disturbance of external dynamic load, rapid mining is more likely to induce rock burst. [ABSTRACT FROM AUTHOR]

Protective layer mining is an important regional measure to reduce pressure and prevent rockburst. Based on the engineering background of frequent mining earthquakes in the 203 working face below the upper protective layer of the Kuangou Coal Mine, PFC was used to simulate the mining process of the protective layer and protected layer, and many factors affecting the pressure relief effect of protective layers were analyzed in detail. The results indicate that the arch force chain structure of the high overlying rock transfers most of the load to the outside of the goaf, and the low-caving broken rock strata hinder the transmission of the remaining load, resulting in a decrease in stress below the protective layer. The uneven collapse of the overlying rock causes the stress relief zone and stress recovery zone under the protective layer. The distance between coal seams and interlayer rock mechanical strength led to rockburst beneath the protective layer. Increasing the mining height and range of the protective layer enhances the pressure-relief effect, while increasing the coal seam spacing and rock strength reduces the pressure-relief effect of the protective layer. This study enriches the theory of protective layer mining. It provides a basis for the evaluation of pressure-relief effect and safe mining of working face under similar mining and geological conditions. Highlights: The PFC simulation method of protective layer mining is proposed. The pressure-relief mechanism and corresponding rockburst causes of protective layer mining are studied. The influencing factors are studied and corresponding explanations are given in pressure-relief effect of protective layer mining. The principle of roadway layout in protected layer mining is given. [ABSTRACT FROM AUTHOR]

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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. [ABSTRACT FROM AUTHOR]