Abstract:Objectives To investigate the relationship between the development of a quasi-granular arch structure within a loose layer and surface damage,this study examines the structural evolution of the quasi-granular arch and the overburden movement behavior under a thick loose layer.Methods Taking the 2303 working face of Hemei No.5 Mine as the engineering background,PFC particle flow numerical simulation,on-site drilling leakage observation combined with microseismic exploration,and nonlinear regression analy-sis were comprehensively employed.Results The results indicate that,with the advance of the working face,the overall failure mode of the bedrock exhibits a positive trapezoidal shape.Particles at the bottom of the loose layer are disturbed by bedrock bending and subsidence,forming an arch-shaped bonded failure zone.During particle movement and compaction,a quasi-granular arch structure with a certain bearing ca-pacity is gradually formed.The fissure field,stress field,and displacement field show distinct stage-dependent evolution characteristics.The development height of the water-conducting fracture zone experi-ences four stages:rapid growth,slow growth,abrupt increase,and stabilization.The stress field evolution undergoes three stages:development of the bedrock pressure arch,coupled evolution of the pressure arch and quasi-granular arch,and independent development of the quasi-granular arch.Surface subsidence evolves through four stages:slow subsidence,accelerated subsidence,rapid subsidence,and stabilization,corresponding well to the staged evolution of the three fields.Surface subsidence measurements based on a surface monitoring circle show that the subsidence curve is approximately V-shaped.After mining,step-like cracks appear in the central area of the goaf,while tensile cracks develop on both sides,consistent with field investigation results.Field observations indicate that the height of the water-conducting fracture zone reaches 147.46 m,with a fracture-to-mining ratio of 16.49,while the caving zone height is 57.46 m,with a caving-to-mining ratio of 6.43.Based on regression theory,a multivariate nonlinear prediction model for maximum surface subsidence under thick loose layers is established,effectively reducing prediction errors caused by single-factor consideration and achieving a fitting coefficient of 0.98.Conclusions The proposed results and methods provide a useful reference for coal mine safety production and surface protection under similar geological conditions.

Abstract:Objectives In response to the frequent accidents such as gas leakage and environmental pollu-tion caused by natural gas pipelines in gas-coal overlap zones,this study takes pipelines in four typical overlapping mining areas of S Mine,X2 Mine,H Mine,and M Mine in the Daniudi Gas Field as ex-amples to conduct a risk assessment of natural gas pipelines in the regions.Methods By analyzing the causes of pipeline damage,a risk assessment system was constructed,consisting of three first-level indica-tors(pipeline factors,environmental factors,and personnel management factors)and seventeen second-level indicators(such as pipeline subsidence and pipeline corrosion).Subsequently,a risk assessment model was built based on extension theory.The BP neural network was employed to calculate the weights of the indicators.The correlation degrees of the indicators were then calculated using distance correlation func-tions and similarity functions,respectively,followed by a quantitative risk analysis of the pipelines.Re-sults The results show that the risk correlation degrees for pipelines in the four areas,calculated based on the distance correlation function,were 0.002,0.034,0.053,and 0.019,respectively.The corresponding risk similarities calculated based on the similarity function were 0.054,0.510,0.560,and 0.500,respec-tively.The corresponding risk levels for the four areas were"high risk,""moderate risk,""moderate risk,"and"low risk."Conclusions Both calculation methods based on the distance correlation function and the similarity function are applicable in the extension evaluation model.The assessment results are largely con-sistent,jointly verifying the reliability of the evaluation outcomes.

Abstract:Objectives To elucidate the evolution characteristics of mining-induced surrounding-rock damage and surface subsidence under complex geological conditions involving extremely thick water-bearing uncon-solidated layers,and to investigate the effects of unconsolidated-layer thickness,bedrock thickness,and aquifer dewatering on surface subsidence.Methods The Huaibei mining area,characterized by extremely thick water-bearing unconsolidated layers,was selected as the study region.Numerical simulation was con-ducted using FLAC3D software to implement a stress-damage-seepage coupled mathematical model for coal mining beneath unconsolidated layers.The evolution of mining-induced surrounding-rock damage,perme-ability coefficient,and surface subsidence in both mined and unmined areas was analyzed.Results After mining,the surrounding-rock damage zone and high-permeability zone exhibited a saddle-shaped distribu-tion.Increasing unconsolidated-layer thickness led to the expansion and interconnection of primary fractures in the weathered zone,forming preferential seepage pathways that guided groundwater infiltration from the fourth aquifer.This induced aquifer dewatering consolidation and intensified surface subsidence,exhibiting a growth pattern characterized by an initial rapid increase followed by gradual stabilization,with the central area of the subsidence basin becoming relatively flat.Increasing bedrock thickness caused the evolutionary height of the damage and high-permeability zones to first rise and then decline.Thin bedrock resulted in greater surface subsidence and more extensive damage zones,while thick bedrock effectively restrained damage evolution in the weathered zone,reduced or prevented infiltration from the fourth aquifer,and miti-gated compressive subsidence caused by aquifer dewatering.Conclusions Both unconsolidated-layer thick-ness and bedrock thickness significantly influence the distribution of mining-induced surrounding-rock dam-age and permeability evolution.The extremely thick unconsolidated layer acts as an overburden load on the bedrock,while bedrock thickness regulates the extent of damage evolution by controlling the bearing capac-ity of key strata.The weathered zone in direct contact with the fourth aquifer is a critical factor contributing to aquifer dewatering.These findings provide a theoretical basis for water inrush prevention and surface sub-sidence control in coal mining beneath extremely thick water-bearing unconsolidated layers.

Abstract:Objectives To improve resource recovery rates and alleviate the tension in mining succession,the experimental study on the mechanical properties of foam filling materials for the roof of ultra-high gob-side entry in thick coal seams was conducted.Methods To verify the adaptability of the foam material,ex-perimental methods were employed to study the influence of factors such as height-to-width ratio and load-ing rate on its mechanical characteristics.Results The results indicate that,under a constant loading rate,the bearing capacity of the foam material gradually decreases as the height-to-width ratio increases.Con-versely,with a constant height-to-width ratio,the bearing capacity gradually increases with the loading rate.Based on the complete stress-strain curve,the loading process of the foam material can be divided into three stages:In the initial loading stage,the working resistance of the specimen increases sharply,representing a brief linear-elastic stage.As loading continues,stress increases slowly with strain,entering the plastic platform stage where the curve shows a clear linear characteristic.Stage III is the densification stage,where specimens with larger height-to-width ratios exhibit a significant reduction in bearing capac-ity,displaying characteristics of softening and structural failure.Graphite polystyrene foam material exhibits notable elastic aftereffect.After the initial loading,it rebounds by 50%to 80%,demonstrating strong deformation-yielding capacity and rebound-bearing performance.Conclusions Combining the engineering geological conditions of the ultra-high roadway in Yunding Coal Mine and based on the porous foam model and the theory of effective extension of anchor cables,it is concluded that foam material with a height-to-width ratio of 0.5 to 1.0 possesses strong bearing capacity and high compression modulus.It exhibits charac-teristics of strong initial deformation resistance and structural stability,enabling it to adapt to large defor-mations of the roadway roof and sides and meet the need to eliminate gas accumulation space.It is thus the preferred material for filling the roof of ultra-high roadways.

Abstract:Objectives To address the problems of severe surrounding rock stress concentration,complicated advance support procedures with poor effectiveness,and slow working-face advancement in advanced mining roadways of thick coal seams under three-soft conditions,a study on synergistic surrounding rock control technology integrating roof cutting and support was conducted.Methods Taking the belt transportation road-way of the 12203 working face in Zhaojiazhai Coal Mine as the engineering background,numerical simula-tions combined with field industrial tests were employed to analyze the factors influencing surrounding rock stress concentration in the advanced roadway section.The optimal roof-cutting scheme and advanced anchor-cable reinforcement support scheme were investigated.Results The results indicate that the superposition of the advanced and lateral abutment pressures of the working face with the lateral abutment pressure induced by roadway excavation has a significant influence on surrounding rock stress concentration in the advanced support zone.Numerical simulations demonstrate that roof cutting can effectively alleviate stress concentra-tion by transferring surrounding rock stress,modifying the load distribution of the mining area,and reliev-ing pressure.The optimal roof-cutting parameters were determined as a cutting angle of 10° and a cutting depth of 18 m,and the corresponding advanced anchor-cable reinforcement support scheme was estab-lished.Field application of the synergistic"roof cutting-support"technology in the belt transportation road-way of the 12203 working face shows that stress concentration in the shallow surrounding rock of the ad-vanced zone was effectively mitigated,advance passive support was eliminated,and safe roadway operation was ensured.Conclusions The synergistic surrounding rock control technology integrating roof cutting and support can effectively relieve stress and reinforce support,ensuring the safety and stability of advanced mining roadways while enabling safe and efficient working-face advance.This technology provides a valu-able technical reference for surrounding rock control in similar thick coal seam mining roadways under three-soft conditions.

Abstract:Objectives To reveal the failure evolution mechanism of the end-area hanging roof of a coal min-ing face and to achieve rational mine pressure control,this study investigates the static fracturing failure characteristics of the hanging roof and determines the optimal drilling spacing.Methods A numerical model of end-area hanging roof failure was established using the Continuous-Discontinuous Element Method(CDEM)to analyze the static fracturing failure characteristics.Based on the geological conditions of the 5-2 coal seam in a representative mine,a numerical model of the strata in the end area of the working face was constructed to investigate the deformation and fracture processes of the hanging roof before and after static fracturing during excavation,thereby revealing the failure evolution mechanism under the action of static cracking agents.Results The results indicate that the expansion pressure of the static cracking agent is positively correlated with the tensile strength of the hanging roof.When the tensile strength is lower than 8 MPa and the expansion pressure is≥30 MPa,optimal fracturing performance is achieved.The optimal drilling spacing for hanging roof fracturing is 1.25 m.Before static fracturing,the periodic caving interval of the end-area hanging roof ranges from 37.5 m to 22.5 m,with an average interval of 27.5 m.After static fracturing,the periodic caving interval is reduced to 17 m to 12 m.Field measurements of support resis-tance and roof displacement further verify the feasibility of the static fracturing technique.Conclusions By regulating the caving interval and fracture sequence of the hanging roof,static fracturing technology can ef-fectively mitigate safety hazards induced by sudden mine pressure variations.The findings provide theoreti-cal guidance for the engineering application of static fracturing technology in hard roof control.

Abstract:Objectives In engineering practice,rocks are often subjected to cyclic loading and unloading conditions,making it essential to investigate the failure precursors in red sandstone under such conditions.Methods Mechanical monitoring tests with acoustic emission(AE)measurement were conducted on red sandstone under cyclic loading and unloading.The k-means clustering algorithm was employed to perform cluster analysis on AE characteristic parameters,aiming to investigate the rock damage and fracture pat-terns.Results The results indicate that acoustic emission(AE)events occur predominantly during the cy-clic loading phase,while very few events are recorded during the unloading phase,marking an"intermit-tent period".As the number of cycles increases,the peak stress gradually rises,and the number of AE events surges abruptly,entering an"active period".At the initial stage of cyclic loading and unloading,the AE b-value is the highest.As the number of cycles increases,the AE b-value gradually decreases and stabilizes within the range of 0.75～1.25.Before the rock enters the fracture stage,the AE b-value shows a sharp drop.However,a similar sharp decline in the AE b-value is also observed at the beginning of each loading cycle,even though the rock does not fracture.This suggests that using a sharp drop in the AE b-value as a precursor indicator for rock failure has limitations,and it is necessary to combine other phenom-ena to study rock fracture precursors.Based on k-means clustering,the AE signals during the rock damage and fracture process are classified into three categories:cluster 1,cluster 2,and cluster 3.Each cluster corresponds to different damage stages.In the late stage of cyclic loading and unloading,the energy of AE events in cluster 3 increases sharply,reflecting a significant enlargement of fracture scale in the sandstone during the final loading stage until failure.This can serve as a characteristic precursor to rock failure.Con-clusions Cluster 3 acoustic emission signals sporadically appear during the elastic deformation stage of the rock.As cycling accumulates,these signals become continuously and densely concentrated prior to the com-plete failure stage of the rock.Integrating the analysis of the AE b-value with the cluster-based signals can enhance the accuracy of predicting rock damage and fracture.

Abstract:Objectives To ensure the safe withdrawal of hydraulic supports during the final mining stage of a fully mechanized top-coal caving face under a thick coal seam with a hard roof,this study investigates the optimization of deep-hole pre-splitting blasting parameters and evaluates the pressure-relief effectiveness of roof cutting in the overlying strata.Methods Taking the 12314 working face in the Wangjialing mining area as the engineering background,a structural mechanics model and a numerical model of a large-space stope during the final mining stage were established.A dynamic coupling analytical method between the overlying strata load and the support bearing capacity was developed,and a stress-blocking positioning method for roof cutting and pressure relief was proposed.Results During the final mining stage,the cantilever length and height of the roof structure were identified as the dominant factors controlling support loading.Increas-ing the roof-cutting height and reducing the distance between the cutting position and the stop line signifi-cantly enhanced the stress-blocking effect.The stress-blocking efficiency gradually stabilized with increas-ing roof-cutting height,with a critical height of 35 m.In contrast,the stress-blocking effect weakened as the distance between the roof-cutting position and the stop line increased,with a critical distance of 60 m.Based on the nonlinear relationship between roof-cutting parameters and pressure-relief indicators,a predic-tion model for roof-cutting pressure-relief effectiveness was developed using surface fitting,achieving a pre-diction error of only 3.32%.Under the optimal roof-cutting parameters for the 12314 working face(roof cut-ting conducted 35 m ahead of the stop line with a cutting height of 30 m),the peak advanced support stress was reduced by 25.93%,the maximum roof subsidence decreased by 14.10%,and the peak stresses on the left and right sides were reduced by 9.83%and 5.50%,respectively.Conclusions The designed com-bination of roof-cutting parameters for the final mining stage exhibits a pronounced stress-blocking and pressure-relief effect,effectively ensuring the safe withdrawal of hydraulic support.The proposed prediction model enables rapid calculation and visualization of multiple pressure-relief indicators,facilitating the de-sign and optimization of key parameters for deep-hole pre-splitting blasting.Furthermore,the research framework of"theoretical analysis-numerical simulation-prediction analysis-on-site verification"and the stan-dardized procedure for roof-cutting parameter design provide a valuable reference for similar engineering ap-plications.

Abstract:Objectives To investigate the deformation behavior and permeability evolution of raw coal under different gas pressures,and to clarify the mechanical-seepage coupling characteristics of coal under gas-bearing conditions.Methods Triaxial compression-seepage experiments were conducted on raw coal samples under constant confining pressure and varying gas pressures.The permeability evolution during loading was analyzed.Based on the Mohr-Coulomb strength criterion,the cohesion and internal friction angle of coal un-der different gas pressures were calculated.A six-stage seepage evolution model was established to quantita-tively characterize the relationship between permeability and volumetric strain rate.Results Under constant confining pressure,coal permeability exhibits a distinct V-shaped variation with increasing axial stress.With increasing confining pressure,both the initial permeability and the peak-state permeability decrease exponentially.Increasing gas pressure weakens the mechanical parameters of coal,including cohesion and internal friction angle.According to the monitoring results of axial stress and gas flow during axial loading,the seepage process can be divided into six stages:stable adsorption,attenuation,gradual increase,abrupt increase,steady increase,and rapid surge.After the peak stress,gas permeability shows a linear re-lationship with the volumetric strain rate.Conclusions This study reveals the dynamic evolution of coal per-meability under the coupled effects of axial stress,confining pressure,and gas pressure,and clarifies the influence of gas pressure on coal mechanical properties and the permeability-deformation relationship.The results provide important theoretical support for determining pressure-relief zones,designing gas extraction sequences,and optimizing borehole parameters in protective seam mining.

Abstract:Objectives To address the challenge of large amounts of coal resources constrained beneath ur-ban buildings,which restrict the sustainable development of coal mines,Chengjiao Coal Mine was selected as a test site to investigate rational backfilling mining technologies and process designs for coal seams be-neath urban buildings.Methods Using engineering analogy,scheme comparison,and theoretical calcula-tion methods,and based on the principle that backfilling mining beneath buildings must prevent surface subsidence,four mining schemes were proposed:fully mechanized mining with cemented paste backfill-ing,new solid backfilling mining,super-high-water backfilling mining,and overburden isolation grouting backfill mining.The strata control effects of different schemes were analyzed,and the optimal scheme was determined as fully mechanized mining with cemented paste backfilling.Support height was determined ac-cording to coal seam thickness,and the working resistance of the support was calculated based on resis-tance to lateral paste pressure and control of roof subsidence at the filling face.A coordinated"isolation-backfilling-coal mining"process was designed to eliminate mining-filling interference.Based on the roles of the backfilling body during the solidification and coal mining stages,the early strength,stability safety fac-tor,and later-stage strength of the backfilling body were determined.Mix proportion experiments and me-chanical property tests of paste backfilling materials were conducted.Results The minimum and maximum structural height of the paste backfilling support are 2.0 m and 3.8 m,respectively,with an extension ratio of 1.9.The working resistance of the support exceeds 8 250 kN.The compressive strength of the paste back-filling body is not less than 0.06 MPa during the solidification stage and not less than 0.16 MPa during the coal mining stage.With a strength stability safety factor of 2.5,the uniaxial compressive strength of the paste backfilling body after 28 days of curing is not less than 2.97 MPa.Economical mix proportions that satisfy strength requirements were obtained for fly ash paste and gangue-fly ash paste backfilling materials.Conclusions The rational backfilling mining technology and process design for coal seams beneath urban buildings applied at Chengjiao Coal Mines can provide valuable references for similar mines in China.

Abstract:Objectives To investigate the overburden migration behavior and surface subsidence characteris-tics induced by top-coal caving mining in the extremely weakly cemented Cretaceous strata of Eastern Inner Mongolia,a numerical simulation study was carried out using a typical working face of the Wujianfang coal-field as the engineering background.Methods A particle flow numerical model was first established based on PFC2D 5.0,and the rationality of model parameters was verified by comparing the simulated surface subsidence with in-situ measured data.The overburden migration characteristics were then analyzed by inte-grating force chain diagrams,principal stress distributions,and displacement fields.Finally,a sensitivity analysis was conducted to evaluate the effects of burial depth,mining height,and advancing speed on over-burden migration.Results The results indicate that the overlying strata generally experience three stages:formation and rapid dissipation of a caving arch,gradual formation of a subsidence basin,and periodic ex-pansion of the subsidence basin.The peak range of the front abutment pressure is 8.19～11.86 MPa,and the maximum deviatoric stress reaches 10.76 MPa at a mining distance of 200 m.Although the overall stress level is relatively low,it is significantly higher than the mechanical strength of the rock strata.The maximum surface subsidence values at the end of the three stages are 0.95 m,6.67 m,and 8.85 m,re-spectively,which are notably larger than those observed in western Jurassic coal mines.Sensitivity analysis shows that the R-values of mining height,burial depth,and advancing speed are 5.37,2.76,and 2.01,re-spectively,indicating that mining height has the greatest influence on overburden migration.Conclusions The findings provide theoretical guidance for the prevention and control of mining-induced hazards in ex-tremely weakly cemented Cretaceous strata in Eastern Inner Mongolia.

Abstract:Objectives The influence of pressure relief borehole diameter on the long-term stability of pres-sure relief roadways is investigated through graded loading creep tests on coal rock specimens with different borehole diameters.Methods First,creep characteristics and long-term stability of coal rock are compara-tively analyzed for different pressure relief borehole diameters.Second,Najar damage theory is introduced to analyze damage characteristics of coal rock during creep stages from an energy perspective.Third,con-sidering the hardening effect of viscosity coefficient as a time-dependent non-stationary parameter,a param-eter nonlinear constitutive equation is proposed.Finally,the steady-state creep rate inflection point method is employed to determine the long-term strength of coal rock with different borehole diameters.Results The results indicate that the pressure relief borehole diameter significantly influences creep characteristics,creep damage and long-term strength of coal rock.During creep,the pressure relief borehole diameter ex-hibits a distinct size effect on both creep strain and initial creep rate,with larger boreholes resulting in greater creep strain and higher initial creep rate.Throughout the creep process,damage gradually increases while damage rate decreases and eventually stabilizes,with damage showing a positive correlation with borehole diameter.The derived parameter-nonlinear Maxwell creep equation,incorporating with the viscos-ity coefficient,accurately describes the creep process.Long-term strength demonstrates a negative correla-tion with borehole diameter.Conclusions The findings contribute to maintaining long-term stability of pres-sure relief roadways and provide theoretical insights for rockburst prediction and mine safety.

Abstract:Objectives To investigate the mechanism of roof cutting and pressure relief in roadways with im-mediately overlying thick-hard roof strata adjacent to working faces,and to determine the rational width of small coal pillars for gob-side entry driving.Methods Taking the return air roadway of the 15107 working face in the No.15 coal seam of Tongbao Coal Mine as the study object,theoretical calculations,numerical simulations,and on-site industrial tests were conducted to analyze the stress distribution characteristics of surrounding rock during gob-side entry driving with small coal pillars under conditions of roof cutting and pressure relief in immediately overlying thick-hard roof strata.Results The results show that a roof-cutting height of 21 m in the transportation roadway of the 15106 working face can effectively cut off the overlying thick-hard roof strata,reducing the lateral cantilever length of the roof.The theoretically determined reason-able width of the small coal pillar for gob-side entry driving ranges from 4.807 to 5.225 m.After roof cut-ting,the surrounding rock stress in the solid coal on the right side of the goaf decreases significantly,re-sulting in a more moderate stress environment for the gob-side entry.Roof cutting also effectively reduces the peak vertical stress within the small coal pillar.Numerical simulation results for coal pillar widths of 3,5,7,and 9 m indicate that increasing the pillar width from 3 m to 5 m leads to an increase of up to 166.31%in the peak vertical stress on the roof surface of the coal pillar,whereas increasing the width from 5 m to 9 m results in peak stress variations of less than 27.32%.Conclusions Field tests demonstrate that leaving a 5-m-wide small coal pillar and implementing appropriate support measures for the 15107 return air roadway can limit the surrounding rock deformation to less than 200 mm under the influence of working-face mining.Deep-hole pre-splitting blasting for roof cutting and pressure relief exhibits a significant control effect,satisfying the safety production requirements of the 15107 working face and providing a reference for coal mine production under similar geological conditions.

Abstract:Objectives To address the problem of low early strength of fly ash paste filling material,this study investigates the mechanism and effectiveness of synergistic modification using high-power ultrasound and alkali activation.Methods Single-factor experiments were conducted to analyze the effects of ultrasound time,ultrasound power,Na2SiO3 concentration,and NaOH concentration on the 1 d,3 d,and 7 d peak strengths of fly ash paste filling material.Based on a three-factor,five-level central composite design(CCD)within the response surface methodology,the synergistic modification parameters of ultrasound and alkali activation were optimized.Furthermore,the microstructural changes after modification and the syner-gistic modification mechanism were systematically investigated.Results The results indicate that the peak strength of fly ash paste filling material first increases and then decreases with increasing ultrasound time,ultrasound power,Na2SiO3 concentration,and NaOH concentration.The degree of influence of the four fac-tors on peak strength follows the order:Na₂SiO₃ concentration>ultrasound time>ultrasound power>NaOH concentration.The optimal experimental conditions for the synergistic modification were an ultra-sound time of 28.62 min,an ultrasound power of 700 W,and a Na2SiO3 concentration of 0.57 mol·L-1.The synergistic action of high-power ultrasound and alkali activation significantly enhanced the peak strength of the fly ash paste filling material,in which alkali activation played a dominant role,while high-power ultra-sound exerted a strengthening effect.Conclusions The synergistic effect of high-power ultrasound and alkali activation further accelerates the hydration reaction rate of fly ash paste filling material,increases the for-mation of hydration products,reduces material porosity,and consequently improves its mechanical proper-ties.

Abstract:Objectives The distribution characteristics of the deep in-situ stress field in the Pingdingshan mining area is to be investigated.Methods Based on an analysis of geological structure,a new method for delineating regional tectonic units was proposed using the main controlling structures—the Guodishan Fault and the Likou Syncline—as boundaries.Then,an analysis curve for the lateral pressure ratio was derived from the Hoek-Brown strength criterion to quantitatively assess the influence of regional geological tectonic on deep in-situ stress field.Results The results show that the deep in-situ stress field in each regional unit,influenced by different geological structures,exhibits different distribution patterns.The stress state in the western unit is predominantly normal faulting.Three kinds of stress states coexist in the central unit.The eastern unit is characterized mainly by the strike-slip and reverse faulting stress state.The vertical stress in-creases linearly with depth in all regional units.The mean value,variation range,and stress gradient of the maximum horizontal principal stress all increase from the west to the east.Similarly,the range and ampli-tude of the lateral pressure ratio increase from the west to the east,indicating that horizontal tectonic activ-ity intensifies from the western to the eastern part ofthe mining area.The analysis of regional geological structures further reveals that the distribution of deep in-situ stress field follows the regional tectonic evolu-tion law.The analysis of regional geological structures further reveals that the distribution of the deep in-situ stress field follows the regional tectonic evolution pattern.Specifically,the influence coefficient of re-gional tectonic stress field is highest in the eastern region(1.07),followed by the central region(0.55)and lowest in the western region.Conclusions Therefore,the evaluation method for regional geological structures based on the lateral pressure ratio analysis curve is applicable to the study of deep in-situ stress fields in the Pingdingshan Mining Area.

Abstract:Objectives To address impact-induced instability of gob-side roadways beneath immediately over-lying thick,hard roof in deep mining faces of the Yanzhou mining area,the transport roadway of the 18306 working face in Jisan Coal Mine is taken as the engineering background to investigate the underlying disaster mechanisms and propose effective control technologies.Methods A comprehensive approach inte-grating field investigation,theoretical analysis,numerical simulation,and in-situ testing is employed.A cantilever-beam mechanical model of the immediately overlying thick,hard roof is established to systemati-cally analyze its mining-induced load-bearing behavior and the evolution characteristics of energy accumula-tion.The influences of roof properties on the catastrophic instability of gob-side roadways are revealed,and the definition and evolutionary process of dynamic instability under thick,hard roof conditions are clarified.On this basis,a layered collaborative control system integrating variable-diameter borehole sealing and coupled roof anchoring–grouting reinforcement is proposed and implemented.Results 1.The accumulation and distribution of mining-induced energy in thick,hard roof strata are significantly influenced by their in-herent geological and mechanical properties.Higher roof strength and greater thickness correspond to lower peak bending energy and weaker energy concentration,whereas longer roof suspension lengths and greater overlying loads result in higher peak bending energy accumulation and a more extensive high-stress zone.2.Dynamic instability of deep roadways essentially results from rapid energy release and structural degrada-tion of surrounding rock under strong dynamic disturbances induced by thick,hard roof fracturing,typi-cally manifested as coal ejection and support system failure.3.A layered collaborative control strategy com-bining pressure relief through variable-diameter borehole sealing and roof anchoring–grouting reinforce-ment is developed,forming a targeted protection system for immediately overlying thick,hard roofs.Con-clusions Field tests demonstrate that the proposed technology effectively ensures the stability of surround-ing rock in gob-side roadways beneath immediately overlying thick,hard roofs,providing a valuable refer-ence for support design in similar roadway conditions.

Abstract:Objectives With the increase in coal seam mining depth and intensity,high-energy mine tremors frequently occur during the excavation of rockburst-prone coal mine roadways.Therefore,it is necessary to study the layout parameters of pressure relief boreholes on both sides of rockburst-prone roadways and their effectiveness in pressure relief and rockburst prevention.Methods Using the excavation of a mining road-way in the pillar area of the 16082 bottom-slice fully mechanized caving face in Qianqiu Coal Mine as the engineering background,an orthogonal experimental design was employed to establish various pressure re-lief borehole layout schemes with different depths,diameters,and spacings.The stress fields of the sur-rounding rock for each pressure relief scheme were numerically calculated to quantitatively study the pres-sure relief and rockburst prevention effectiveness,and the results were validated through field industrial tri-als.Results The results show that:(1)Using the transfer distance of the high-stress zone D as a quantita-tive evaluation index,when D≥1.42,the abutment pressure curve at the roadway sides transitions from a"single-peak"to a"double-peak"pattern,indicating sufficient pressure relief in the surrounding rock and a significant effect on pressure relief and rockburst prevention.(2)Range-variance analysis revealed that the influencing factors on the degree of high-stress zone transfer are,in order of significance:borehole depth,borehole diameter,and borehole spacing.Borehole depth primarily governs the transfer distance of the high-stress zone.The maximum transfer distance is positively correlated with borehole depth,while the average transfer distance is positively correlated with borehole diameter.When the borehole depth is constant,the average transfer distance shows a linear negative correlation with borehole spacing.Conclusions Field indus-trial trials demonstrated that after optimizing the pressure relief borehole parameters,the energy of mining tremors during roadway excavation was predominantly at 10² J,with fewer high-energy tremors of 10³ J and above.The convergence of the roof to floor and the average drilling cuttings volume from observation bore-holes decreased by 43%and 30%,respectively.Therefore,a quantitative threshold for the transfer dis-tance of the high-stress zone and an optimization system for pressure relief borehole parameters are pro-posed,providing theoretical guidance for efficient pressure relief in rockburst-prone roadways.

Abstract:Objectives The prediction accuracy of surface movement and deformation under complex mining conditions is to be improved,and the surface movement and deformation induced by inclined irregular working face mining is to be accurately predicted.Methods Based on a unit influence function that consid-ers variations in mining depth,a calculation method for the mining depth of arbitrary units in inclined coal seams was proposed,and a subsidence prediction model suitable for inclined irregular working faces was established.The model applies Green's theorem to convert subsidence area integrals into boundary line in-tegrals,derives the corresponding line integral prediction formula,and provides a numerical solution method.On this basis,an integrated prediction program was developed,achieving full-process computation from parameter input and determination of integration intervals to line integral solutions.The program en-ables visualization of prediction results,including surface subsidence,tilt,curvature,horizontal move-ment,and horizontal deformation,and further evaluates and analyzes the mining-induced damage levels.The proposed model was validated and analyzed through case studies of inclined working faces in actual mining areas.Results The results show that the proposed prediction model and integrated program can effec-tively address the accurate prediction of surface subsidence and classification of mining-induced damage un-der complex mining conditions.The prediction results conform to the fundamental laws of mining subsid-ence,with a relative error in subsidence prediction of less than 10%,thereby meeting engineering accu-racy requirements.Conclusions The proposed model is reliable and practical,with high prediction accu-racy,and can provide a theoretical basis and technical support for the mining design optimization,precise surface damage assessment and protection strategies in inclined irregular working face mining.

Abstract:Objectives The use of filling materials for roadside support in gob-side entry retaining under fully mechanized top-coal caving mining is often associated with large material transportation volumes,com-plicated construction processes,low early strength,and high costs.Therefore,there is a strong need to de-velop a high-strength,high-stiffness,lightweight,and modular roadside support structure.Methods The bearing performance of grouted steel pipe piers was investigated through theoretical analysis,laboratory ex-periments,numerical simulations,and engineering application.Steel pipe specimens with a length of 500 mm and a diameter of 159 mm,fabricated with different wall thicknesses and grouted with slurries of vari-ous water–cement ratios,were subjected to axial loading tests and numerical simulations to identify the key factors influencing bearing capacity and to determine reasonable design parameters.Based on structural optimization,steel pipe piers with a length of 1 200 mm and a diameter of 159 mm were manufactured and grouted with a slurry at a water–cement ratio of 0.4,followed by axial loading tests and numerical simula-tions.Furthermore,taking a gob-side entry retaining project in a fully mechanized top-coal caving working face as the engineering background,the bearing performance of field-used grouted steel pipe piers(length 3 000 mm,outer diameter 299 mm,wall thickness 8 mm,water-cement ratio 0.4)was analyzed using the limit equilibrium method and numerical simulations.Results The results indicate that a steel pipe wall thickness of 8 mm combined with a water–cement ratio of 0.4 provides an optimal ultimate bearing capac-ity for the grouted steel pipe piers.The axial ultimate bearing capacity of the piers applied on site exceeded 4 800 kN,which is in excellent agreement with theoretical predictions.Conclusions By optimizing the steel pipe wall thickness and slurry mix proportion,grouted steel pipe piers exhibit high bearing capacity while simplifying the construction process and reducing equipment requirements and material costs.Engi-neering practice demonstrates that when grouted steel pipe piers are applied for gob-side entry retaining un-der fully mechanized top-coal caving mining,the maximum convergence between the roof and floor is 78 cm,and the maximum convergence of the roadway ribs is 95 cm after stabilization,meeting the engineer-ing requirements for roadway deformation control.

Abstract:Objectives The stability of reverse faults and the overlying strata movement pattern under mining influence is investigated to prevent fault activation and instability triggered by mining activities.Methods Through a combination of engineering case studies,theoretical analysis,and numerical simulations,the fault activation mechanism under mining disturbance is systematically analyzed.The susceptibility of fault activation during hanging-wall and foot-wall mining operation is quantitatively assessed and the correspond-ing failure modes and fracture development patterns in the overlying strata are compared.Results The re-sults show that:(1)During hanging wall mining of the reverse fault,both normal and shear displacement vectors of the overlying rock joints approach 0,indicating minimal mining impact on strata movement and fault activation.Although stress concentration is observed near the fault but no slip displacement occurs.(2)During foot-wall mining,the maximum overlying strata settlement is 0.16 m higher than during hanging-wall mining,with greater fracture development height.The normal and shear displacement vectors of the overlying rock joints reach 0.4 m and 1.3 m,respectively.Under the influence of mining,the fault exhibits both stress concentration and activation-induced slip instability.(3)Using the activation risk index Q to evaluate the fault activation risk,foot-wall mining is found to demonstrate higher susceptibility to fault slip activation.(4)The stability criteria for faults during hanging-wall and foot-wall mining are determined to be T/F>0.087 and T/F>11.43,respectively,indicating greater instability risk during foot-wall mining.Significant differences in fault stability are observed between the two mining scenarios.(5)For foot-wall mining,a protective coal pillars with a horizontal distance exceeding 117 m between the working face and the fault is required.Conclusions These results provide critical insights for optimizing coal resource recov-ery while ensuring mining safety in areas affected by reverse fault structures.