In order to study the failure characteristics of overlying strata during the mining of thick coal seams under the conditions of shallow burial depth, thick loose layers, and thin bedrock, this study was based on the geological conditions of the 13101 working face in the Shaping Coal Mine. By comprehensively applying methods such as theoretical analysis, numerical simulation, and on-site measurement, a study is conducted on the overlying strata failure law during the working face mining process. The results show that: (1)-Through the observation of flushing fluid loss, the development height of the water-conducting fracture zone in the working face is 88.6-92.2 m, with a fracture-mining ratio of 7.9-8.2; the results of numerical simulation and theoretical calculation are both relatively close to the measured results. (2) With the mining of the working face, the surface movement and subsidence trend evolves from a ″V″ shape to a ″U″ shape. Affected by the liquidity index of the thick loose layers, the surface subsidence lags behind the upward development process of the overburden failure height; (3) There is a Boltzmann functional relationship between the overburden failure height and the mining distance, and its evolution can be divided into four stages: initiation, rapid increase, slow increase, and stability. The first three stages mainly occur in the thin bedrock, while the stability stage appears in the thick losse layers; (4) The failure of overlying strata exhibits a "two-belt" destruction pattern, in which the thin bedrock acts as a support, while the thick loose layer mainly provides high static load and has an inhibitory effect on the degree of overburden destruction.

With the growing depth of underground mining, issues surrounding solid waste storage and the effective use of mine water have become pivotal to achieving sustainable mining practices. The complex ionic composition of mine water impacts the performance of traditional cemented tailings backfill (CTB) materials. Gold mine tailings, combined with cement, were repurposed as the cornerstone raw materials in this investigation. Solutions with identical target concentration gradients for Cl−, SO42−, and HCO3− were prepared separately using NaCl, Na2SO4, and NaHCO3, respectively, with each salt dosed to achieve the desired anion concentration. These solutions served as mixing water for preparing samples with tailings and cement. Strength, energy dissipation characteristics, and microstructure of CTB were investigated by single-axis compression test, XRD, and SEM-EDS analysis. Experimental results demonstrate that adding three reagents—NaCl, Na2SO4, and NaHCO3 (covering Cl−, SO42−, and HCO3− ions, respectively)—at appropriate concentrations enhances mechanical properties. At their optimum concentrations, these salts increased the compressive strength of CTB by approximately 30%, reaching ~4 MPa. However, further increases in salt concentration produced inconsistent strength responses, with bicarbonate-containing mixtures exhibiting the most pronounced strength reduction. These effects are primarily attributed to competition between the introduced anions and cement hydration reactions, which alters the pore structure and consequently the density and strength of the hardened matrix. Incorporating different ion-covering backfill at appropriate concentrations enhances mechanical strength. These findings provide new opportunities for CTB mix design and mine water utilization. However, as this study considered only single-ion systems, further investigation is needed to elucidate the combined effects of multiple ions present in actual mine water.

Mining coal seams with shallow, thick, and hard roofs often results in extensive roof suspension. This issue poses significant challenges regarding stratum control and mitigation of strong mining pressure, especially within the confined working space of a mining face. This study focuses on the 13101 working face of Shengfu Coal Mine. Through field observations, theoretical analysis, and numerical simulations, the characteristics of support resistance and microseismic activity were investigated. This research elucidates the mechanism behind the strong mining pressure driven by the structural coupling and synergistic breakage of two key strata, highlighting how their interaction dictates weighting intensity. A small-aperture hydraulic fracturing technology, specifically designed for inter-support spaces, was developed. The results indicate that the working face exhibits alternating “minor weighting” and “major weighting” events. Minor weighting occurs at an average interval of 12.38 m with a dynamic load factor of 1.14, while major weighting occurs at 41.07 m with a factor of 1.56. The roof structure was found to form a combination of an “inclined stepped rock beam” and a “voussoir beam.” Field applications demonstrate that the proposed technology reduces the major weighting interval by 41.46% and total microseismic energy release by 35.01%. This study provides a theoretical and technical basis for preventing roof disasters under similar geological conditions.

The fracture movement and structural instability of hard roof strata in large-mining-height faces are fundamental triggers for rock burst occurrences in mining panels. Elucidating the correlation between the fracturing of roof strata and mining-induced seismic events is therefore of significant importance for the scientific prevention and control of rock burst. The high-energy seismic events and rock burst manifestations at the Yingpanhao Coal Mine, operated by a subsidiary of Yankuang Energy Group in Ordos City, are investigated with the 2215 gob-side face employed as a case study. Focusing on the concentration of large-energy events ahead of the working face and along the gob-side edge, the research integrates theoretical analysis, numerical simulation, and field measurement. By applying an elastoplastic plate structure ultimate load analysis method, the fracture characteristics of roof strata and the geometric dimensions of plastic hinged blocks under conditions of a large-mining-height gob-side face were determined. A Fish program was developed to simulate the evolution of dissipated energy during strata fracturing, enabling the characterization of the distribution and dynamic evolution of dissipated energy in roof strata as the face advances. The main fracture trace distribution and the structural morphology of broken blocks in the roof strata of large-mining-height faces are identified, thereby revealing the structural configuration of roof fracturing. Through this, the mechanism behind the concentration of high-energy events in the front abutment pressure zone and the gob-side area is elucidated. The findings indicate that: The ultimate load-bearing capacity of the roof strata decreases exponentially with increasing face advance distance until the initial and periodic fracturing intervals are reached. Moreover, the roof in the large-mining-height gob-side face exhibits eccentric fracture characteristics, with the main fracture location shifting toward the gob-side edge and the rear of the goaled area. An overhanging structure is formed in the roof strata above the front abutment and the gob-side area. The interaction of successive fracture traces results in the formation of an approximately “L”-shaped roof block, whose rupture and movement are identified as the primary cause of high-energy microseismic events in these zones. In thick and hard multi-layer strata, the implementation of high-density advanced destressing boreholes during the mining cycle of the two gate roads effectively disrupts the combined structural effect and holistic movement behavior of multi-layered “L”-shaped roof blocks. This measure has proven effective in reducing the occurrence of high-energy microseismic events in large-mining-height faces with hard roof strata at the Yingpanhao Mine, which provide valuable references for safe extraction of high-quality coal resources under similar geological and mining conditions.

In this study, the thick, hard roof group of the 20103 working face in the Dahaize Coal Mine is used as the research object. UDEC discrete element simulations were conducted to examine fracture-induced energy release characteristics under unweakened conditions and single-layer hydraulic fracturing at different burial depths. This study clarifies the evolution of elastic strain energy accumulation and dissipation in thick, hard strata; reveals the post-fracturing energy migration mechanisms of different horizons; and establishes the correspondence between fracturing horizon and energy-weakening effectiveness. The results show the following: (1) Without fracturing, the energy accumulation within 0–30 m ahead of the face differs markedly among key strata. KS1 exhibits stable and relatively uniform elastic strain energy accumulation; KS2 shows a rapid increase at an advance distance of 180–300 m; and KS3, being farther from the coal seam, accumulates less energy than KS2. (2) After individually fracturing KS1, KS2, or KS3, the fractured layer loses its bearing and energy storage capacity, and the adjacent upper key stratum becomes the new load-bearing and energy accumulation layer. The reductions in elastic strain energy accumulation and release vary significantly by fracturing horizon: KS1 decreases by 10.70% and 11.63%, KS2 decreases by 18.73%, and KS3 decreases by 20.83% and 9.10%, respectively. Among the single-layer fracturing schemes, fracturing KS2 provides the most effective weakening of fracture-induced energy release in the mining-disturbed roof strata.

In order to study the loading state of the coal body in front of the heading face of the outburst coal seam and the nonlinear mechanical transfer path of the in-situ coal and rock mass under static and dynamic loads, a mathematical model of the transversely isotropic layered combined coal and rock mass was established by using the Hamilton mechanical system, and the symplectic space-time subdomain method was used to solve the interlayer dynamic mechanical transfer path of the coal and rock mass. This study investigates the unloading-reloading process of excavation-induced stress waves and its effects on the coal-rock mass ahead of the tunneling face. The results show central stress concentration with attenuation along the short and long axes, and an "X"-shaped shear failure along the diagonal. Plastic loading waves overtaking elastic unloading waves induce delamination at 0.02, 0.03 and 0.06 s, while radial deformation causes additional delamination at 0.004 and 0.028 s. Analysis based on the Mohr-Coulomb criterion and strain energy density reveals a "V"-shaped energy pit, consistent with experimentally observed cross-shaped fractures dominated by axial propagation and long-axis extension. This study conclusions can provide a theoretical basis for predicting and preventing coal-rock instability.

The destabilizing damage of rock structures in coal beds engineering is greatly influenced by the bearing rupture features and energy evolution laws of rock–coal assemblages with varying height ratios. In this study, we used PFC3D to create rock–coal assemblages with rock–coal height ratios of 2:8, 4:6, 6:4, and 8:2. Uniaxial compression simulation was then performed, revealing the expansion properties and damage crack dispersion pattern at various bearing phases. The dispersion and migration law of cemented strain energy zoning; the size and location of the destructive energy level and its spatiotemporal evolution characteristics; and the impact of height ratio on the load-bearing characteristics, crack extension, and evolution of multiple energies (strain, destructive, and kinetic energies) were all clarified with the aid of a self-developed destructive energy and strain energy capture and tracking Fish program. The findings indicate that the assemblage’s elasticity modulus and compressive strength slightly increase as the height ratio increases, that the assemblage’s cracks begin in the coal body, and that the number of crack bands inside the coal body increases as the height ratio increases. Also, the phenomenon of crack bands penetrating the rock through the interface between the coal and rock becomes increasingly apparent. The total number of cracks, including both tensile and shear cracks, decreases as the height ratio increases. Among these, tensile cracks are consistently more abundant than shear cracks, and the proportion between the two types remains relatively stable regardless of changes in the height ratio. The acoustic emission ringing counts of the assemblage were not synchronized with the development of bearing stress, and the ringing counts started to increase from the yield stage and reached a peak at the damage stage (0.8σc) after the peak of bearing stress. The larger the rock–coal height ratio, the smaller the peak and the earlier the timing of its appearance. The main body of strain energy accumulation was transferred from the coal body to the rock body when the height ratio exceeded 1.5. The peak values of the assemblage’s strain energy, destructive energy, and kinetic energy curves decreased as the height ratio increased, particularly the energy amplitude of the largest destructive energy event. In order to prevent and mitigate engineering disasters during deep mining of coal resources, the research findings could serve as a helpful reference for the destabilizing properties of rock–coal assemblages.

In response to the common issue of instability and failure of weak interlayers under load in geotechnical engineering, a discrete element simulation study plan consisting of 25 schemes was designed, with the inclination and thickness of the interlayer as variables. Based on the test results from a universal testing machine, the weak interlayer and adjacent rock mass simulation parameters were calibrated. A Fish program was developed to monitor the evolution of cracks, the number of tension and shear cracks, block elastic strain energy, and tension and shear strain energy throughout the uniaxial compression process. This study reveals the influence of the inclination and thickness of the interlayer on the evolution of the "main rupture crack morphology—changes in the number of tension and shear cracks—energy accumulation and dissipation" under uniaxial compression conditions, as well as the interrelationship among these three factors. The results show that (1) when the interlayer thickness is the same, as interlayer inclination increases from 0° to 60°, the strength of the specimen decreases by 38%-40%. Elastic strain energy is positively correlated with the peak strength of the specimen. When interlayer inclination angles are 0°, 15°, 30°, and 45°, the number of shear cracks and shear strain energy stored in the contact do not change significantly. However, when the interlayer inclination is 60°, compared to the specimen with the interlayer inclination of 45°, the number of shear cracks in the specimen decreases by 24.01% to 48.28%, and shear strain energy decreases by 35.06% to 50.35%. (2) When the interlayer inclination is the same, as the interlayer thickness increases from 4 to 12mm, the strength of the specimen decreases by 6.49% to 22.02%. The specimen’s total tension cracks increase by 38.1% to 143.52%, while tension strain energy decreases by 13.04% to 40%. The number of shear cracks and shear strain energy fluctuates, with increases being positive and decreases being negative, with ranges of -7.65% to 26.02% and -11.33% to 24.55%, respectively. (3) Under the influence of different interlayer thicknesses and inclinations, the number of shear fractures and their energies are significantly higher than those of tension fractures, and shear failure is the main reason for the failure and instability of weakly cemented interlayered rock mass. The research results can provide a reference for predicting and evaluating rock and soil mass failure modes and designing reinforcement schemes for projects with weak interlayers.

In response to the challenges of controlling surrounding rock deformation in gob-side entry driving towards the advancing working face, a systematic study on the stability of the headgate# 15107 and coal pillar section was conducted, using a combination of theoretical analysis, numerical simulation, and field testing. First, based on the theory of internal and external stress fields, the range of the internal stress field was determined to be 9.83~11.43 m, and combined with the limit equilibrium theory, the most reasonable width of the narrow coal pillar was found to be 6 m. Secondly, the stability of the surrounding rock and coal pillars of the headgate# 15107 under different coal pillar widths during roadway excavation and working face mining was simulated, respectively. The simulation results show that during the head-on mining and driving period, when the coal pillar width is 4 m or 5 m, the plastic zone in the coal pillar is completely damaged and loses its bearing capacity; when the coal pillar width is 6 m, an elastic zone appears in the coal pillar, and the area of the elastic zone increases with the increase in the coal pillar width. During the excavation along the goaf, when the coal pillar width is 4, 5, 6, 8, or 10 m, the stress curve inside the coal pillar shows a single-peak distribution, and the stress peak of the coal pillar increases with the increase in the coal pillar width, with the stress peaks being 7.66, 9.74, 12.32, 16.02, and 27.05 MPa, respectively. When the coal pillar width is 25 m, the stress curve inside the coal pillar shows a double-peak distribution. During the advancement of the 15107 working face, the stress peaks corresponding to the 4, 5, 6, 8, 10, and 25 m coal pillars are 29.8, 27.5, 26.8, 27.2, 33.7, and 24.3 MPa, respectively. Throughout the entire simulation process, when the coal pillar width is 6 m, the coal pillar has good bearing capacity and a low degree of stress concentration. Finally, based on this, the support scheme for the headgate# 15107 was optimized, and industrial experiments were conducted. Field testing showed that a 6 m narrow coal pillar for roadway protection and an optimized roadway support can effectively control the deformation of the surrounding rock of the roadway.

Based on considering the stress state distribution and potential failure surface of the specimen during uniaxial compression, the drilling parameters (layout, spacing, position) are set. Thoroughly understanding the influence of different drilling parameters on the pressure relief effect is conducive to reducing the occurrence of coal mine rock burst accidents. Through laboratory tests and numerical simulation tests under different drilling parameters, the influence laws of mechanical parameters, failure characteristics, AE characteristic parameters and energy evolution of specimens under different drilling parameters were studied. The pressure relief effect under different drilling parameters was evaluated by using the pressure relief effect evaluation index (X), and the best combination of drilling parameters was obtained. The results show the following: (1) Compared with the intact specimen, the peak strength of the drilling specimen is significantly reduced, and the drilling layout has the greatest influence on the mechanical properties, followed by the drilling spacing and drilling position. (2) Different drilling layouts will form different weak-strength surfaces in the specimen, and lead the expansion and penetration of cracks, resulting in different failure modes of the specimen. The stress distribution inside the specimen will affect the stress concentration around the borehole, finally affect the damage degree of the specimen. (3) Drilling can not only effectively reduce the energy accumulation capacity, but also enhance the degree of energy dissipation. The AE ringing counts and energy of the triangular-drilling specimens are the least. The AE ringing counts and energy decrease first and then increase with the increase in drilling spacing, and are the smallest at three times the drilling diameter. The AE ringing counts and energy increase gradually with the upward movement of the drilling position. (4) The optimal combination of drilling parameters was obtained by the test, and it was triangular-layout drilling, drilling spacing three times the diameter, and the drilling position in the middle of the specimen, and the value of the pressure relief effect evaluation index (X) was 65.41. The research results can provide some reference for the selection and optimization of drilling pressure relief parameters in rock burst mines.

Macrostructures and inclusions are both vital for slabs because the quality of slabs is largely affected by them. However, the relationship between macrostructures and inclusions in the thickness direction of the slab is still unclear. Hence, in this paper, the relationship between macrostructures and inclusions was revealed by laboratory experiments and theoretical calculations. The laboratory experiments included carbon and sulfur content testing, direct reading spectroscopy, scanning electron microscopy, and automatic inclusion scanning. The experimental results showed that the distribution of macrostructures was symmetrical from the inner and outer arc to the center. From the edge to the center of the slab, the variation in macrostructures was columnar crystal zone (CZ)→columnar-to-equiaxed transition (CET)→equiaxed crystal zone (EZ). Furthermore, the content of sulfur and manganese first decreased and then increased from the inner arc to the outer arc. The number density and area fraction of MnS inclusions in different macrostructures were CZ > CET > EZ. The average size of MnS in different macrostructures was CZ > EZ > CET. Moreover, the morphology of MnS inclusions was ellipse and rod in CZ, irregular dendrite in CET, and multilateral in EZ. Additionally, theoretical calculation results showed the maximum precipitation and initial precipitation temperature of MnS inclusions in different macrostructures were CZ > EZ > CET. Meanwhile, the theoretical precipitation radius of MnS inclusions in different macrostructures was CZ > EZ > CET.

High in-situ stress and mining disturbance will lead to the deformation and instability of deep rock mass. Under the action of high in-situ stress, deep rock mass enters the strength limit neighborhood, and the rock mass entering the strength limit neighborhood is very sensitive to mining disturbance. Research on micro-damage test of rock rheological disturbance effect is the key to solve the problem. In order to meet the test requirements in the research process, a micro-damage test system for rock rheological disturbance effect was developed. The test system optimized the host of the tester by upgrading the disturbance loading device, and expanded the micro-damage test function of rock rheological disturbance effect. The test results show that: ① The performance of rock rheological disturbance effect micro-damage test system is reliable, which is suitable for the related research of rock rheological disturbance effect micro-damage test; ② When the rock enters the strength limit neighborhood, the growth rate of large-sized pores with a pore size greater than 1 μm is much larger than the growth rate and pore compaction amplitude of small-sized pores with a pore size less than 1 μm, resulting in irreversible damage inside the rock; ③ Under different confining pressure rheological disturbance conditions, when the specimen is outside the strength limit neighborhood, the confining pressure has a certain inhibitory effect on the expansion and development of the pores of the specimen, and the compaction amplitude of the pores inside the specimen is greater than its expansion amplitude. With the increase of confining pressure, when entering the strength limit neighborhood, the existence of confining pressure accelerates the expansion and development of pores to a certain extent, and the internal pores of the specimen are dominated by expansion and development; ④ The microscopic criterion for the rock to enter the strength limit neighborhood is determined, that is, when the proportion of large-sized pores with pore size greater than 1μm increases significantly under the condition of rheological disturbance, it indicates that the rock has entered the strength limit neighborhood.

In order to disclose the multi-physical field characteristics of the deep coal seam mining process and their dynamic evolution legislation, based on the “rock-coal-rock” model, during the mining process, the stress field, displacement field, energy field, and plastic zone evolution process are all simulated using FLAC3D6.0. The findings show that stress in the original rock is redistributed as a result of coal seam mining, creating a pressure relief zone in the middle of the goaf and advanced support pressure in the front part of the working face. The roof falls following the termination of coal seam mining. The collapsed blocks fill the middle of the goaf, playing a supporting role. The floor bulges as a new supporting pressure zone forms and builds up high elasticity. The stress reduction zone shifts from a rectangular to an inner circular distribution and an outer square as the working face’s mining distance increases and the range of the fracture field expands accordingly. In addition, a complete model was constructed to verify the correctness of the “rock-coal-rock” model. The stress, displacement, and energy curves of the overlying strata at a distance of 12 m from the bottom of the coal seam in the middle of the goaf obtained by the two methods were basically consistent. Ultimately, the findings of the numerical simulation were compared with the advanced support pressure data that were acquired on-site and they were good. This work can provide a reference for the safe mining of deep coal seams.

Confined aquifers widely exist in the strata of the Ordos mining area. Water drainage before mining is an effective measure to avoid water inrush disasters caused by the connectivity between mining-induced roof fractures and aquifers. However, rock burst disasters occur frequently in the mining process of many water drainage working faces. The statistics show that the surrounding rock of nine water drainage working faces in five mines has suffered different degrees of rock burst, which seriously restricts the safe, efficient production and sustainable development of ten-million-ton modern mines in China. Based on this, taking the 22,106 longwall working face (LW22106) of the Shilawusu Mine, Ordos, as the engineering background, this paper investigates the occurrence mechanism of water drainage on rock burst in the roof water-rich area by using theoretical analysis, similar material simulation, microseismic measurement and other methods. The main conclusions are as follows: (1) After the drainage of the water-rich area, the pressure relief zone, pressurized zone and pressure stabilization zone are formed in sequence from the center to both sides. The width of the pressure relief zone is consistent with that of the water-rich area, and the width of the pressurized zone is about 35 m on one side. (2) When the mining is passing the pressurized zone, the coal rock mass is under the joint influence of gravity stress, mining disturbance stress and drainage transfer stress. The superimposed stress generally exceeds 2.5 times the uniaxial compressive strength (UCS) of the coal rock mass, and the maximum reaches 3.24 times, far exceeding the critical value of rock burst (1.5 times UCS), which is the main reason to induce rock burst. (3) The dynamic change in the rock burst risk areas before and after drainage in water-rich area no. 4 has been predicted, and the number of risk areas increases from 4 before drainage to 13 after drainage. Since the stress superposition effect differs due to different mining speeds, it is proposed that the dynamic regulation of mining speed in the front and at the back of the drainage area is an effective and efficient method for rock burst prevention and control. The research results will provide a theoretical basis and technical support for the prevention and control in the roof water-rich area of deep mines.

Mine seismic events are an inevitable dynamic phenomenon occurring in deep mines. A scientific and rational method is needed to evaluate and understand mine seismicity and its induced disasters. In the Ordos mining area of North China, multiple groups of thick hard-bedded sandstone formations commonly exist in the overlying strata of Jurassic coal seams. In recent years, frequent mine seismic events in many large mines of Ordos have resulted in suspended or limited production, which seriously threatens the safe and efficient operation of 10-million-ton modern mines in China. Therefore, taking the frequent occurrence of mine seismic events in the mining process of goaf working face with a multi-layer thick hard roof in Ordos mine as the research background, this study investigated the mechanism and prevention of mine seismic in goaf working face with the methods of case study, theoretical analysis and field monitoring. The following conclusions are made: when the goaf working face is mined, an “advanced and lateral” L-form roof forms under the coupled influence of the lateral suspension plate formed above the upper working face and the roof of the working face. Due to the common influence from “advanced and lateral” L-form roof activation, the gradually breaking multi-layer thick hard roof, thick hard roof group bending and prying effects, in addition to excessively fast or uneven mining speed, mine seismic events will occur frequently when the exceedance warning index (EWI) is breeched. On this basis, coordinated blasting to break the roof along two roadways and within the working face is put forward as a measure with the purpose of preventing and controlling mine seismic events, and a robust effect on mine seismic reduction and disaster prevention is obtained in field application. The research results can serve as a reference for the development and application of mine seismic mechanism and blasting vibration reduction technology on the working face where there is a multi-layer thick hard roof, thereby supporting a strategy of promoting the resource development and energy security of deep mines.

To investigate the evolution law of inclusions in 42CrMo-S steel, this paper samples and analyzes the steel during its refining process as well as the head and tail billets. An oxygen and nitrogen analyzer, a scanning electron microscope (SEM) equipped with energy-dispersive X-ray spectrometry (EDS), and an ASPEX automatic inclusion scanning electron microscope are employed to analyze the cleanliness level of the molten steel in the refining stage and the head and tail billets. The results demonstrate that the total oxygen content at the end of LF slagging is 10.2 ppm, indicating that the refining slag has an excellent deoxygenation effect. During the RH refining process, the total oxygen content of the molten steel diminishes to less than 10 ppm and reaches 6.3 ppm at end-RH. The nitrogen content in the molten steel gradually increases during the smelting process and attains 65 ppm at end-RH. Upon arrival at LF, pure Al2O3 plays the role of the primary inclusions in the molten steel. Afterwards, the pure Al2O3 inclusions transform into Mg-Al spinel-type inclusions, Al2O3-MgO-CaO inclusions, and Al2O3-CaO inclusions. The number of CaS-type inclusions in the steel reaches the maximum after feeding the S wire. In the RH refining stage, the percentage of inclusions with a size less than 5 μm is maintained above 90%. Finally, the cleanliness level of the head and tail billets (the start and end of a casting sequence) is analyzed, and it is recommended that the cut scrap length for the head billet is 0.3 m and the reasonable cutting scrap length for the tail billet is 1 m.

针对陕蒙地区深部矿区工作面隔离煤柱宽度设计不合理导致冲击地压和矿震频发的现状,采用案例调研、理论分析和现场监测等方法,对隔离煤柱区冲击地压和矿震的发生机理及隔离煤柱合理宽度进行了研究.以近年来陕蒙深部矿区工作面在隔离煤柱区发生的3起典型冲击地压和矿震事件为工程背景,根据动力显现特征和诱发机理不同,将其分为煤柱局部破坏型冲击地压、煤柱整体失稳型冲击地压和煤柱区厚硬砂岩组破断型矿震3类.分别建立了采空区侧向支承压力估算模型、采空区转移应力估算模型、煤柱承载应力估算模型和关键层挠度弯曲变形力学模型,揭示了不同宽度隔离煤柱诱发局部冲击、整体冲击和矿震的机理.研究结果表明:2101工作面和2201工作面间隔离煤柱不发生局部冲击的宽度为5～6 m或不小于128 m;不发整体冲击失稳的宽度为不小于138 m;不发生矿震的宽度为498 m.在此基础上探讨了不同宽度隔离煤柱对冲击地压和矿震的影响,提出了基于冲击地压-矿震协同控制的合理隔离煤柱宽度设计方法,以期为陕蒙深部矿区相似条件工作面隔离煤柱宽度设计提供参考.

Aiming at the phenomenon that many dynamics occur frequently in working face or roadway when the first mining face of mine A, 2-2upper201 is driven to the edge of water-rich area, the effects on original rock stress by draining water-rich area and rockburst risks in two different conditions of drainage and no-drainage are studied through some methods such as theoretical analysis and numerical analysis, based on working face 2-2upper201A(the second working face of Mine A). The findings are as follows: ①The drainage in water-rich area can damage aquifer, resulting in stress decrease inside damaged area and stress concentration in the edge; ②Considering the effect by draining water-rich area, rockbust scope and degree in damaged area will change. In detail, the range below water-rich with risk will shrink and degree of risk will decline, while the edge and outside of water-rich area will enlarge as risk rises. The field monitoring indicates that in the process of mining, stress increments on measure spots rank as the order of from large to small as follows: edge > outside > inside. When working face is driven to the edge of water-rich area, many drastic dynamics like coal blasting, wall carving will occur on the wall, which suggests that draining water-rich area has a great impact on rockburst risk of working face. The research result has very important guiding significance to rockburst prevention in water-rich area.

Evaporation parameter of solute atoms segregated on surfaces is quantified and its physical meaning is revealed to be an integrated effect of segregation, diffusion, and evaporation. Evaporation activation energy is formulated and its minimum can be used to estimate the binding energy between solute and solvent atoms. Evaporation activation energy of Mg from the surface of Al-0.8 wt%Mg alloy is determined to be very close to the binding energy between Al atom and Mg atom, indicating segregated Mg is adsorbed on the surface. Excellent agreement between theoretical calculations and experimental results provides a quantificational approach to uncover the evaporation mechanism. (C) 2019 Elsevier B.V. All rights reserved.