Abstract:To test anchor combination components and metal mesh in case of support failure in complex and difficult roadway support process,a static loading test device for anchor combination components and metal mesh was developed,and a series of vertical load tests for support components was carried out.The mechanical response characteristics of support components under vertical load have been obtained.Firstly,a mechanical performance test bench for anchor support combination components and metal mesh static load was developed,and then a set of boundary fixing methods for metal mesh and anchor rod support combination components was proposed.The developed testing system can restore the stress state of underground metal mesh and anchor support combination components more realistically.During the load-bearing process of reinforced beams,the load changes are mainly divided into three stages:slow rise stage,constant speed rise stage,and instantaneous failure stage.The load variation of W-shaped steel strip also goes through three stages:bearing stage,long hole sliding stage,and long hole tearing stage.By increasing the diameter of the steel bar support beam and the thickness of the W-shaped steel strip,the support strength and stiffness of the support system can be effectively improved.Among them,the weak strength areas,such as the welding parts of the steel bar support beam bracket and the long hole ends of the W-shaped steel strip are the key factors affecting the support strength and stiffness.The load displacement curves of different types of metal mesh support systems under vertical loads are mainly divided into three stages:load transfer stage,mesh deformation stage,and structural failure stage;Steel mesh has high stiffness support characteristics,diamond mesh has high strength support characteristics,and the continuous displacement of mesh wires under load is the main factor weakening the support strength and stiffness of the warp and weft mesh support system.The shear and tensile breakage of mesh wires at the tray position is the main factor causing the failure of the metal mesh support system.

Abstract:To improve the supporting capability of the hinged canopy of the four-leg shield powered support and to address the common perception of inadequate support provided by these canopies,a rod system model for the four-leg shield powered support was developed.This model was employed to determine the bearing capacity zone of the powered support based on the loading conditions.It yields the maximum load that the powered support can withstand and identifies its specific action location on the canopy.Two methods of the hinged canopy support force method and the hinged cylinder diameter methodfor solving this problem are presented.The simplified resolution method transforms the external load acting on the canopy into various distribution forms,including quadratic polynomial,triangular,trapezoidal and rectangular shapes,depending on the position of the external load,as described by Taylor's formula.By integrating the hinged structure of the canopy with the assumption of continuous load distribution above it,the magnitude and point of action of the external load on the hinged canopy are determined using the principle of force synthesis.This analysis allows for the derivation of the required diameter for the hinged cylinder based on static equilibrium conditions and specific formulas for solving the problem are provided.The support forces at the front end of the hinged canopy is compared with those of the overall top canopy under identical working resistance conditions.The findings indicate that a larger diameter hinged cylinder significantly enhances the support capacity of the hinged canopy,establishing that the support capacity of a well-designed hinged canopy-powered support surpasses that of an equivalent overall canopy-powered support under the same conditions.Using the mining practices of Shandong Energy Zhengtong Coal Industry Company's four-leg shield-powered support as the primary application context,a continuous load distribution model for the hinged canopy of the ZF20000/26/48 four-leg shield-powered support was established.Based on the continuous load distribution conditions above the canopy,the size and location of the external load on the hinged canopy were determined using the principles of force synthesis and decomposition.Furthermore,the theoretical diameter of the hinged cylinder was derived from static equilibrium conditions.Engineering practice demonstrates that the improved hinged canopy structure is rational,exhibiting significant front-end support force and pronounced support effects.It has been widely applied in various mining faces without encountering issues such as insufficient front-end support force leading to roof collapse or coal layer spalling,thereby validating the accuracy of the theory and effectively meeting the support needs of the mining face.

Abstract:Ground uplift caused by groundwater rebound in closed coal mining areas threats infrastructure safety.We proposes a novel method for predicting surface uplift that does not rely on in-situ measurements of groundwater levels or geological/mining data.The method is based on the characteristic'S'-type growth pattern of surface uplift revealed by the principle of effective stress.The Weibull function and the Kalman filter are integrated for prediction model development.The Weibull function captures the long-term main trend,while the Kalman filter dynamically assimilates time-series InSAR monitoring data to account for actual fluctuations.An empirical study conducted at the closed Guanshan and Taiji mines in Beipiao City,Liaoning Province,is sued to validate the proposed model.The prediction shows an average deviation of only-1.2 mm and an average standard deviation of 2.9 mm when compared to leveling survey data,achieving an overall prediction accuracy of approximately 3 mm.By combining historical InSAR deformation data with the proposed model,this approach overcomes the dependency of traditional methods on field monitoring data,enabling accurate prediction of the surface uplift process in closed mining areas.This research not only provides a new methodological perspective for analyzing the mechanisms of mining-induced surface deformation but also offers efficient and reliable technical support directly applicable to land reclamation planning,infrastructure risk prevention and control in mining regions.

Abstract:The stability and failure modes of tectonic coal under disturbance differ significantly from those of original coal.To investigate the energy evolution characteristics and disparities during the fracturing process of deep tectonic coal,the mesoscopic parameters,macroscopic mechanical properties,energy distribution of tectonic coal and microseismic characteristics in tectonic coal-bearing strata were analyzed through various loading paths combined with acoustic emission monitoring,computed tomography(CT scan)reconstruction and microseismic monitoring.The result shows distinct microstructural differences exist between tectonic and origin coal,manifested as poor surface regularity,abundant debris accumulation and stepped fractures in tectonic coal.These microstructural characteristics lead to divergent macroscopic mechanical behaviors,with linear correlations between mesoscopic and macroscopic parameters.Tectonic coal exhibits substantially different characteristic stress stages under loading when compared to origin coal,showing shorter elastic phases and stable crack propagation periods.While primary fractures induce varying degradation intensities,similar crack development patterns emerge beyond critical stress thresholds.Tectonic coal demonstrates more dispersed fracturing patterns with greater susceptibility to cyclic loading effects.Regardless of loading paths,tectonic coal tends to develop large-scale cracks with more disordered spatial distributions.Shear-dominated failures with tortuous crack paths prevail in tectonic coal,requiring higher energy consumption.The elastic energy dissipation ratio decreases during cyclic loading-unloading processes in fractured coal,accompanied by increased proportions of released energy and maximum strength reduction.Microseismic monitoring shows fewer low-energy events in primary coal zones when compared to tectonic coal areas.Tectonic coal-bearing strata exhibit higher microseismic frequency with gradual energy release patterns,while origin coal strata demonstrate more sudden seismic events,particularly requiring attention in fractured coal zones.The findings interpret field microseismic data and help early warning systems development for tectonic coal-bearing strata hazards.

Abstract:To address the challenges of global climate change,China has proposed the"dual carbon"strategic goal of"carbon peaking and neutrality".Deep underground engineering,including deep underground energy exploitation,geological carbon dioxide sequestration,deep underground energy storage and the reuse of abandoned mines,provides important techniques for achieving the"dual carbon"goal.Knowledge about the friction-permeability coupling evolution of fractures and faults,both of which are referred to as rock discontinuities here,is of great significance for the safety and efficiency of deep underground engineering.Here,we present a state-of-the-art review on the coupling mechanisms and control techniques of the friction-permeability coupling of rock discontinuities.First,we briefly introduce the key roles of rock discontinuities in contemporary deep underground engineering.Second,basic theories related to the friction and permeability of rock discontinuities as well as their coupling mechanisms are systematically presented.Finally,we analyze the main mechanisms of seismicity occurrence induced by fluid injection and extraction,and elaborate on the mechanisms and challenges associated with the enhancement of reservoir permeability in deep underground energy exploitation and the leakage in deep underground fluid storage,alongside current control techniques.Finally,we propose several important issues to be investigated:the effects of clay on the friction-permeability coupling in rock discontinuities,the friction-permeability coupling in rock discontinuities under constant normal stiffness conditions,the stick-splitting mechanism of hydraulic fractures,etc.Additionally,multi-stage hydraulic fracturing,machine learning,and distributed optical fiber sensing may provide more efficient and intelligent solutions for predicting and controlling the reservoir permeability enhancement in deep underground energy exploitation,leakage risks in deep underground fluid storage,and induced seismic hazards in various deep underground engineering projects.

Abstract:When the mining face approaches the coal pillar,dynamic instability occurs in front coal pillar.To explore the influence mechanism of mining disturbance on dynamic instability of coal pillar,the dynamic compression test of high-stress coal pillar under low-frequency disturbance load was carried out.The results show there is a stress threshold value for the failure of coal samples under high stress(80%of the Unilateral Constrained Compressive Strength(UCCS)),and the threshold value is 15%of UCCS.Subjecting to the stress over the threshold value,the number of disturbance load cycles before the failure of high-stress coal samples decreases exponentially with the increase of amplitude,and the increase of amplitude significantly deteriorates the bearing capacity of coal samples.Also,the irreversible strain,dissipated energy,secant modulus and damping ratio of coal samples show a three-stage evolution law under disturbance load.In addition,with the increase of amplitude,the cumulative dissipated energy in the process of coal sample disturbance increases linearly with the maximum value of 241.932 kJ/m3.The mean value of secant modulus decreases linearly with the minimum value of 1.101 GPa.However,the mean value of damping ratio increases exponentially with the maximum damping ratio of 1.593,indicating the increase of amplitude makes the internal deterioration of coal sample more serious.The large increase of AE energy in the initial disturbance loading stage of coal sample can be used as an advanced criterion to judge whether the coal sample is damaged,and the second surge of cumulative ring count can be used as the precursor information of coal damage beyond the threshold value.With the increase of amplitude,the b value of AE in the whole disturbance loading process on coal decreases linearly with the minimum value of 1.369.In other words,the larger the amplitude,the more sudden and severe the damage of coal sample.With the increase of amplitude,the fracture mode of coal samples when subjecting to stress beyond the threshold gradually changes from shear fracture to tensile or mixed fracture.

Abstract:To address the challenges in controlling hydraulic fracture propagation trajectories and stimulated volumes within multi-lithologic coal measure strata,particularly those caused by formation interfaces and interlayer strength variations,a combined finite-discrete element method(FDEM)was used to investigate hydraulic fracture behavior.The results show twelve distinct propagation modes under multifactor coupling conditions,which can be classified into four categories:interface penetration,penetration with interface extension,interface tracking,and interface arrest.A novel hybrid artificial intelligence model integrating BP neural networks with differential evolution(DE)and grey wolf optimization(GWO)algorithms(BP-DEGWO)was developed to predict fracture trajectories and stimulation effectiveness.Key controlling factors,including rock strength contrast coefficient,interface dip angle,interfacial strength,injection rate and perforation angle were identified,and their relative importance under varying in-situ stress conditions was quantified.A further intelligent optimization framework for hydraulic fracturing design in stratified formations was proposed based on the BP-DEGWO model.This approach provides a predictive tool for fracture geometry in heterogeneous coal-bearing strata and a methodological reference for AI-assisted fracturing optimization.The findings are instructive in enhancing stimulation efficiency in multi-lithologic unconventional reservoirs.

Abstract:Taking the 2201 working face of Haiwan Coal Mine No.3 as a case,the overburden collapse patterns during mining beneath open-pit slopes were studied using physical simulation,numerical simulation and theoretical analysis.A structural model of"large cantilever beam and high-position oblique step voussior beam"was established.The mechanism of strong mine pressure in the initial mining stage under the open-pit slope was revealed.The results show that due to the existence of the open-pit slope,the continuity of the stratum is cut off,and the overlying strata and the surface loess layer form a large cantilever beam fracture and subsidence during the initial mining stage of the working face,which destroys the original caving law of the overlying strata and forms the strong mine pressure.When the working face advanced to 160 m,the advanced fracture development depth in loess layers reached 90 m.When the working face advanced to 178 m,the"large cantilever beam and high-position oblique step voussior beam"structure formed,with peak support load exceeding 50000 kN/set,forming the strong mine pressure.By establishing a mechanical model of the"large cantilever beam",the force balance equations of the rock blocks and the support resistance formula were derived,and the mechanism of strong mine pressure caused by large cantilever beam structure was revealed.A hydraulic fracturing layer that eliminates the formation conditions of large cantilever beam was proposed.By using the numerical simulation method of hydraulic fracturing and support resistance monitoring,it is found that eliminating the large cantilever beam structure can avoid the occurrence of strong mine pressure,and has been verified by engineering practice.

Abstract:Various long and narrow passages are inevitably formed for transportation and ventilation in undergroudn resources and energy exploitation.At depth,failures and instabilities of passages have attracted more and more attention.Using numerical simulation of surrounding rock models containing holes can save time and effort,costing relatively little but obtaining more information.However,reliable results depend on whether advanced numerical methods are used.Recently,a parallel computing system for strata movement(StrataKing)has been developed with a large calculation scale(up to 5 million elements),high calculation efficiency and obvious mining characteristics.After researches on the surrounding rock model containing holes with conventional loading(applying constant stress and velocity on the model boundary),further research,including static and dynamic combined loading,static and dynamic mode transformation and critical stress calculations was conducted.The results show that when the static load exists,the higher the amplitude of periodic impact,the greater the degree of fracture and movement of the surrounding rock of the roadway.This leads to difficulty in balance of surrounding rock of the roadway.As a result,the first impact can cause more fractures,which should be related to the fact that fractures caused by the first impact has a certain energy absorption effect.When the quasi-static calculation model is changed toward the dynamic calculation model progressively,which is more consistent with the disaster gestation and occurrence process,the cracking and movement of the surrounding rock of the roadway are more serious,showing the limitation of traditional quasi-static calculation mode.When the impact energy index is in the actual range,the critical stress of the surrounding rock of the square-profiled roadway is larger than the theoretical result(strictly applicable to circular roadways).By comparison,any shape of the roadway can be considered in StrataKing with large-scale computing capabilities,rich features and mining characteristics.

Abstract:Aiming at the problems of high cost and large filling volume in goaf subsidence control through filling methods,this paper proposes a protective mining method for small-scale surface structures based on"artificial key bearing structure plus dynamic grouting".With theoretical analysis,physical and numerical simulation,this study investigates the mechanical characteristics of the key bearing structure,the stability of the key bearing structure under dynamic grouting,the characteristics of surface movement and deformation,and the influence of the key bearing structure parameters on the grouting filling.The results show that the filling volume is approximately piecewise linearly related to the depth of the key bearing structure.The closer the key bearing structure is to the surface,the smaller the grouting filling volume is.The key bearing structure is constructed in the surface soil layer by splitting grouting method.The stability of this structure and the upper surface soil layer can then be maintained by filling with dynamic grouting in the underlying separation layer and coordinating this process with the mining speed.The relationship model between key bearing structure parameters and grouting filling step is established,and the expression of dynamic grouting filling step is derived.It is found that the maximum filling step increases with the increase of tensile strength and thickness of the key bearing structure,and decreases with the increase of the ratio of topsoil load to building load.After dynamic grouting filling,the peak value of surface horizontal deformation above the key bearing structure is 1.59 mm/m,the peak value of inclination is 2.71 mm/m,and the peak value of curvature is 0.033×10-3 m-1,which are reduced by 87.8%,89.8%,and 80.8%,respectively.The surface deformation meets the requirements of the first-level damage grade of the structure and basically no maintenance is required.

Abstract:Borehole instability,wall collapse and fracture misalignment are the primary causes of borehole failure in gas drainage.To investigate the influence of borehole inclination angle on the load-bearing characteristics and stability of surrounding rock,experiemental study was conducted on the mechanical properties and failure characteristics of surrounding rock with boreholes at different angles.Uniaxial compression tests were conduced on rock-like specimens with prefabricated holes at 0°,30°,60°,75°,and 90° to the horizon.A rock rigidity testing machine,acoustic emission(AE)monitoring system,digital image correlation(DIC)strain measurement system and ultra-high-definition borehole inspection camera were used to study the deformation and fracture behavior of the surrounding rock.The results show that the prefabricated holes reduce the peak strength and elastic modulus of the specimens,with the 60° inclined specimen exhibiting the greatest strength reduction(30.48%lower than the intact specimen).The energy surge pattern during loading correlates with the borehole inclination angle,with the 90° specimen showing a secondary energy release of 75.8 V·ms.Tensile failure dominates,accounting for 70.6%at 0° but decreasing to 51.4%at 90°,while shear failure gradually increases with inclination angle.The load-bearing capacity of the borehole wall increases with the inclination angle,and crack initiation inside the hole lags significantly behind the specimen surface,with failure timing coinciding with AE energy surges.As the borehole angle increases,internal stress and crack distribution shift from concentration at both sides of the hole to a more uniform circumferential distribution,reflecting enhanced load-bearing capacity.The findings help understanding the deformation and failure mechanism of surrounding rock of gas drainage borehole and optimizing borehole layout design.

Abstract:To address the challenge of managing surrounding rock deformation in deep rockburst roadways,this study utilizes the 402101 working face of a coal mine as the engineering background.Through a combination of field investigation,theoretical analyses,numerical simulations and industrial tests,the instability characteristics of the roadway,identifies the primary controlling factors was examined,and the stress transfer mechanisms under static-dynamic loading conditions was elucidated.A novel"bolt-U typed steel support-filling(BUF)"synergetic control technology is proposed and subsequently validated through practical engineering applications.The results show that the principal factors contributing to surrounding rock instability include the interaction of high static-dynamic loads,the inadequate impact resistance of support materials and the poor synergy within the support system.The BUF system establishes a composite bearing structure by employing pre-stressed anchor cables(to control crack propagation),the yielding of U-typed steel supports(to facilitate energy conversion),and the uniform energy dissipation provided by the filling layer.In accordance with the theory of surrounding rock structure stratification,the surrounding rock of the roadway is categorized into three layers:the support layer,the anchoring layer and the original rock layer.The compressive stress of the support layer is influenced by the rockburst failure coefficient kb,the dimensions of the roadway,the physical properties of the surrounding rock and the support resistance provided by bolts and cables.An impact risk assessment methodology is developed,focusing on the relationship between the compressive stress and the strength of the support layer.Computational examples reveal that the BUF system can decrease the compressive stress of the support layer by 18.5%to 42.1%and withstand dynamic load disturbances at kb=2.0.Numerical simulation results indicate that in comparison to conventional support systems,the BUF system reduces the plastic zone by 79.1%,decreases the displacement of the roof and sidewalls by nearly 90%,fully activates the passive bearing capacity of the supports,and optimizes the energy field distribution.In practical engineering applications,the maximum displacement of the surrounding rock controlled by the BUF system is 43 mm,and the maximum stress on the supports is 0.34 MPa.The structure remains stable,thereby effectively ensuring safe and efficient mining operations.

Abstract:The successful mining of working face of 10 m ultra-large mining height in Caojiatan Coal Mine marks a new stage for the fully-mechanized mining of extra-thick coal seams.Based on multi-source monitoring data,such as mine pressure,displacement and microseismics,the rules of mine pressure behavior and the occurrence mechanism of strong mine pressure in the 10 m ultra-high mining height working face were studied,and a"superimposed arch-beam"structural model of the ultra-high mining height stope was established.A three-in-one surrounding rock control strategy of"active support and protection+regional pressure relief and weakening+all-round monitoring and early warning"was proposed.The results show that the cyclic breaking of the overlying rock in the 10 m ultra-high mining height working face has the characteristics of strong mine pressure,such as strong dynamic load,fast roof sinking speed,long lasting distance of pressure,and high opening rate of support safety valves.The cyclic breaking of the roof of the working face is 10-25 m,the average dynamic load coefficient of periodic pressure is 1.42,and the average maximum shrinkage of support is 0.63 m.During the pressure period,the proportion of hydraulic support safety valves opening rate was more than 50%is 48.1%.The incoming pressure shows regional aggregation characteristics along the tendency direction,distributed in the range of 75-250 m from the nose.The trend direction shows the characteristics of alternating incoming pressure with"large and small cycles",and the average interval between large cycles is 137.5 m.The overlying rock strata of the ultra-high mining face are in a superimposed"arch-beam"structure,and the thick and hard roof in the middle and upper parts overhangs the roof behind the goaf for a long distance,causing the roof to break in front of the face or bend and sink behind the support,causing strong mine pressure appears in face.Reducing the step distance and continuous distance of periodic weighting,reducing the dynamic load coefficient of weighting support,preventing coal wall spalling and avoiding the crushing of hydraulic support are the core elements to control the stability of surrounding rock in 10 m ultra-large mining height working face.The high initial supporting force characteristics of the hydraulic support effectively control the early subsidence of the roof,the high working resistance characteristics relieve the bearing capacity of the coal wall,inhibit the large-area spalling of the coal wall and the rapid subsidence of the roof during pressurization effectively.The double-layer telescopic beam structure of the hydraulic support is combined with the three-level linkage protection device,which realizes the coordinated control of the unsupported area of the roof in front of the support and the protection operation of the ultra-high coal wall,and effectively overcomes the defects of the insufficient protection coverage of the coal wall of the original split protection device.The 5 m3/min high-flow hydraulic fracturing technology,employing a"one-field-one-strategy"zonal parameter design,effectively weakens the thick hard roof,leading to a significant reduction in weighting intervals and dynamic loading coefficients.The real-time warning system of KJ21 mine pressure achieves dynamic monitoring of support operating conditions and roof fracturing.

Abstract:As a crucial geological structure in coal mine,faults directly affect the stability of mining operations and the occurrence of fault-related disasters.To address the current limitation that mine-scale fault complexity assessments cannot effectively quantify the impact of faults on working face operations,a new method is proposed by integrating transparent geological model with numerical simulation to calculate the three-dimensional(3D)faults complexity at the working face scale.On one hand,based on a transparent geological model of the working face,the spatial curvature features of fault plane are extracted through mathematical analysis using recursive search algorithms and the least squares method.On the other hand,a numerical model of the fault-containing working face is constructed using the Anderson fault model and the Mohr-Coulomb(M-C)failure criterion.This model simulates the regional stress accumulation patterns in the rock mass under mining conditions,reflecting the degree of fault zone activation.Subsequently,the entropy weight method is used to integrate the two types of fault features,compensating for the statistical characteristics of the fault structure.The 3D fractal dimension is then employed as the framework to compute the overall 3D faults complexity of the working face under mining-induced conditions.Taking the 40103 working face in the Dafosi Coal Mine of the Binchang mining area as the engineering background,the 3D complexity of fault was calculated at mining distances of 100,80,60,40,20 and 0 m from the DF5 reverse fault.Combined with fault spatial distribution and microseismic data,the variation trends of fault complexity at different mining positions were analyzed.The results indicate that the overall 3D fault complexity increases as mining approaches the DF5 reverse fault.The complexity changes mainly because the fault enters an unstable state during mining,which propagates both laterally and longitudinally as the mining face advances-particularly when approaching the fault.Additionally,zones with intersecting multiple faults are more strongly affected by mining disturbances than those with a single fault,with fault structures evolving from single to intersecting patterns.Furthermore,high-energy microseismic events are predominantly concentrated in areas of higher fault complexity.The evolution of these events closely mirrors the trend of increasing fault complexity,indicating that the more structurally complex a fault is,the more uneven the spatial distribution of microseismic events and the higher the frequency of high-energy occurrences.This highlights the correlation between changes in regional rock stress and the fault instability state reflected by the 3D faults complexity model under mining conditions.

Abstract:The joints and fissures produced by the surface deterioration of rock slopes in open pit mines are the key factors inducing landslides,and timely and accurate fissure detection is an important prerequisite for mining safety.Aiming at the problems of single fissure morphology generated by traditional data enhancement algorithms and the existing slope fissure monitoring methods being limited to local image analysis,a method for detecting fissuress in rock slopes of large-scale open-pit mines using Copy-Paste algorithm has been proposed,in which the fissure morphology is augmented by the Copy-Paste algorithm,and then a three-phase system workflow is established with sliding window algorithm-U-Net accurate segmentation-panoramic feature stitching to overcome the problems of lost and false detection of fissure features in large-scale scenes.The experimental results show that the proposed scheme results in a significant increase in the diversity of cleft samples,and the algorithm achieves 0.22%and 0.85%improvement in the segmentation IoU metrics when compared to the Mask R-CNN and U-Net models using traditional data enhancement,respectively.The slope project of Luoyang Luanchuan Longyu Molybdenum Mine is used for validation,showing the proposed three-phase system workflow has a positive role in promoting the identification and segmentation effect of large-scale fissures,with the mIoU value significantly increasing by 7.09%compared to that the three-phase system workflow is not used.The proposed scheme provides a new technical approach for disaster warning in steep rock slope engineering.

Abstract:Quantitatively describing the development of mining-induced fractures and the evolution of permeability in the double key strata is crucial for the efficient gas extraction from the gob area and fracture zone.Taking the Dingji Coal Mine in the Huainan mining district as the engineering background,numerical simulations were conducted to study the dynamic evolution of overburden fractures during mining under the double key strata.The fracture density and fracture ratio are quantitatively analyzed using a grid-based image division method.Fractal theory is introduced to describe the propagation and closure processes of mining-induced fractures beneath the deep double key strata.Based on the fracture characteristic parameters,a fractal seepage model for overburden fractures is derived,and an optimized criterion for target layer selection in high-extraction tunnels is proposed.The study results show during the deep coal seam mining under double key strata,the mining-induced fractures develop upward and forward as the working face advances.In the horizontal direction,fractures in the center of the gob area are small and dense,while fractures in the working face and drifts are large and sparse.In the vertical direction,fractures below Key Stratum 2 maintain their integrity for a long period,and the fracture area and fractal dimension are greater than those in other strata,indicating that the key strata play a protective role in fracture distribution.According to the fractal seepage model for mining-induced fractures,the overburden permeability is in the range of 10-14～10-11 m2.The permeability distribution follows an"unimodal"-"terraced"-"saddle-shaped"evolution pattern as the working face progresses.After mining,a narrow low-permeability zone forms above Key Stratum 1,and the permeability below Key Stratum 2 reaches its maximum value,exceeding 1×10-11 m2,demonstrating that the double key strata indirectly control the distribution of overburden permeability.The main factors influencing the gas extraction efficiency in high-extraction tunnels include the gas source,gas channels,and tunnel stability.Considering these factors,a concept of an effective gas extraction coefficient for high-extraction tunnels is quantitatively defined from three aspects:the degree of gas accumulation,the duration of high-permeability zones,and tunnel stability.An optimized criterion for selecting target layers for high-extraction tunnels is proposed,and it is applied in a trial working face,achieving good gas extraction results.The research findings provide valuable reference for the study of mining-induced fracture seepage evolution in deep coal seams under double key strata and guide the layout planning of high-extraction tunnel layers in mines.

Abstract:For water hazard prevention and control as well as water resource protection in typical mining areas with thick and extremely thick coal seams under water-bearing strata,nearly 30 years of practical experience in mining under water-bearing strata were summarized and refined.This has led to the development of a water control(protection)technology model suitable for typical mining areas in China.Based on the occurrence characteristics of water-bearing roof in typical mining areas and the spatial relationship with the main coal seam,five water filling models were proposed,that is,weakly consolidated water sand collapse,high pressure progressive water filling,high-positioned thick water-bearing strata permeation,upward and downward cracks progressive conduction water filling,and lateral large channel strong recharge water filling.Based on the dual control of disturbance range and water filling volume,an integrated theory and technical system for controlled(protecting)water in coal mine roof water-bearing strata were established,including precise detection of water-bearing strata,precise control of overburden damage,controllable reduction of water filling volume,and coordinated monitoring and early warning of hydro logical conditions.Key technologies were proposed,such as precise detection using transient electromagnetic methods considering shutdown time effects,height estimation of high-intensity mining-induced conductive fracture zones in thick and extremely thick coal seams,pre-mining precise pre-reduction of roof water-bearing strata based on drainage capacity constraints,weakening hard main control overburden to guide the height of conductive fracture zones,and hydrological monitoring and early warning involving three parameters of absolute value,change rate and trend of water level in water-bearing strata.For the water control practice of roof water-bearing strata in some typical mining areas in China,corresponding water control mining models have been proposed,achieving sucess in water control for mining under different types of water-bearing strata.The research findings provide scientific reference for coal mining under water-bearing strata,realizing safe,efficient and green mining.

Abstract:Taking the thick overlying roof of the 122104 working face in the Caojiatan Coal Mine as the research background,prefabricated crack models of different heights was prepared and CT-based in-situ uniaxial compression tests were conducted to investigate the mechanical characteristics of specimens with prefabricated cracks of different heights,along with their mechanical damage and degradation processes.The experimental results show that compared with intact specimens,the peak strength,acoustic emission energy,total strain energy,elastic strain energy and post-peak strain energy of the prefabricated crack specimens are significantly reduced.The energy dissipation index of the upper,middle,and lower crack specimens increased by 24.7%,28.5%,and 62.6%,respectively.The strength degradation index increased by 21.9%,28.1%,and 54.6%,respectively,and the impact energy index decreased by 19.9%,20.9%,and 49.5%,respectively.Among them,the lower-crack specimen shows the best pressure relief effect,with the lowest initial kinetic energy upon failure.The crack initiation times for the intact specimen and the specimens with upper,middle and lower prefabricated cracks are 89,44,42,and 39 s,respectively.The failure mode shifts from tensile-splitting failure in the intact specimen to tensile-shear composite failure in the cracked specimens.As the crack height decreases,the initiation and development of cracks occur earlier,with increased microcracks and rock debris at the fracture surfaces.In the lower-crack specimen,the surface porosity,maximum pore area and probability entropy of the fracture increased by 98.51%,317.58%,and 5.38%,respectively.Based on the analysis of failure characteristics of specimens with prefabricated cracks at different heights,the damage degradation mechanism is revealed from the perspective of hindered force chain transmission.The findings is indicative for optimizing the selection of pressure-relief stratification in thick roof areas.

Abstract:To study the influence of non-uniform stress on rock damage,failure and the evolution law of damage zones,the mechanical response characteristics and energy evolution law of sandstone with defects under non-uniform stress were studied.The results show that the stress-strain curve stage in the low-stress area is highly consistent with the overall curve of the specimen,and the residual stress fluctuation characteristics of sandstone are highly consistent with the curves in the high and low stress areas.The higher the overall bearing capacity of sandstone,the greater the peak difference between stress areas.It is also found that in the high-stress area,circular defect sandstone shows low stress and high deformation at the peak,while trapezoidal and rectangular defect sandstone shows high stress and low deformation at the peak.In the low-stress area,the initial deformation field is prone to intersect,and trapezoidal defect sandstone has the largest radial deformation,circular defect sandstone has the largest axial deformation,and rectangular defect sandstone has the smallest.In the high-stress area,the strain curve contracts inward,while it expands outward in the low-stress zone.The acoustic emission ring count of sandstone with rectangular and circular defects shows a first increasing,then decreasing,and final increasing trend,while that of trapezoidal defect sandstone shows a continuous increase and sudden increases at multiple points,showing an overall dense-sparse-dense change pattern.In the initial stage,sandstone with rectangular defect generates more shear cracks,while sandstone with circular and trapezoidal defects generates fewer cracks.With the increase of stress,the number of shear cracks in the three defect shapes of specimens shows an increasing trend,and the failure mode of sandstone is mainly shear failure.In the high-stress area,there are more shear cracks and they are more likely to form coalesced cracks,while in the low-stress area,there are fewer cracks and they are mostly tensile cracks.The efficiency proportion in the high-stress area is higher than that in the low-stress area,and the proportion of elastic deformation energy in the low-stress area is higher than that in the high-stress area,especially in the specimen with circular defect shape.Energy dissipation in high-stress regions shows a steady increase,while growth in low-stress regions slows.The dissipation energy accumulation curve in high-stress zone shows a sequential trend of acceleration,uniform speed and deceleration trend,while the dissipation energy of the low-stress zone shows a sequential trend of acceleration,deceleration and acceleration trend.

Abstract:Large deformation of surrounding rock is a great challenging in soft rock underground engineering.We analyzed the microscopic and pore characteristics of typical surrounding rock using scanning electron microscopy(SEM)and established a mechanical model for high pre-stressed long and short anchor cables in a collaborative support system.By integrating test results of surrounding rock loose zones and calculations of surrounding rock pressure,the support parameters for long and short anchor cables are determined and verified.A suitable support scheme for large deformation sections using high pre-stressed long and short anchor cables is proposed.Additionally,numerical simulation techniques are employed to systematically analyze the large deformation of surrounding rock and the supporting mechanism of long and short anchor cables under tectonic stress.The results reveal that the surrounding rock of the DLII11+080 section of the branch tunnel of Haidong Tunnel is primarily composed of calcareous shale,characterized by developed fractures,high porosity,and a loose structure.The microscopic layered structure contributes to the rock's fragility,making it susceptible to external forces.Single-length short anchor cable support is ineffective in providing adequate reinforcement to the surrounding rock.However,the use of both long and short anchor cables enhances support by leveraging their synergistic effects.The short anchor cable primarily mitigates early plastic deformation,while the long anchor cable provides overall constraints to prevent the expansion of deep displacement.Through numerical simulations,the plastic zone and convergence deformation of surrounding rock are analyzed under varying lateral pressure coefficients and cohesion levels.The study also elucidates the influence of tectonic stress and surrounding rock strength on tunnel deformation.The results indicate that the depth of the vault plastic zone under the optimized long and short anchor cable support is 48.63%lower than without any support and 44.38%lower compared to support using 6-meter short anchor cables alone.The combined support of long and short anchor cables significantly enhances the uniformity of axial force distribution,improves the overall support effect,and stabilizes the surrounding rock.This research provides a scientific foundation and practical guidance for controlling large deformations under tectonic stress in soft rock underground engineering.