| Cite this article as: |
Xingping Lai, Huicong Xu, Pengfei Shan, Qinxin Hu, Weixi Ding, Shangtong Yang, and Zhongming Yan, Research on the mechanism of rockburst induced by mined coal-rock linkage of sharply inclined coal seams, Int. J. Miner. Metall. Mater., 31(2024), No. 5, pp.929-942. https://doi.org/10.1007/s12613-024-2833-8
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Rockbursts are a prominent challenge being faced by mines in China, particularly in deep mining [1]. Notably, China boasts steeply inclined and exceptionally thick coal seam resource reserves, comprising over 30% of the analogous global resource reserves. The Urumqi mining area possesses considerable reserves of sharply inclined low-sulfur, low-ash, and high-calorific coal, amounting to an impressive 3.6 billion tons [2]. It is necessary to highlight that the Urumqi mining area falls within the seismic confines of the Tianshan region. The relentless pursuit of high-intensity and deep mining operations in mines with steeply inclined coal seams within this area brings about a confluence of factors such as large inclination angles, exceptionally thick coal seams, specialized mining methodologies, and intricate geological conditions; thus, the appearance of dynamic disasters are highly probable. In the mining district, the prototypical mine hosting urgently inclined coal seams consistently undergoes intense rockbursts at a relatively shallow mining depth of just over 300 m. Meanwhile, other mines located in the western sector of the mining district experience such dynamic disasters only when the mining depth exceeds 600–800 m [3]. As a result, in-depth exploration of the rockburst mechanisms inherent in such steeply inclined coal seams is of paramount significance, as it holds the key to ensuring the secure and effective development of analogous mines.
Currently, local and foreign scholars have researched the mechanisms and control issues of rockbursts in mining coal-rock. Zhang et al. [4] explained the disaster-causing effects of dynamic and static load linkages resulting from surface subsidence, “stress breakdown,” and rockburst during the mining of deep, thick bedrock and extra-thick coal seams. Qi et al. [5] observed that the overlying rock structure is prone to showing lever-type “prying” characteristics under special conditions, and the resulting linkage effect is prone to a strong dynamic response to the high stress caused by fault activation. The researchers [6–7] reported that high-energy microseismic events of 1 × 105 J are more likely to occur in mining areas where the maximum horizontal principal stress and lateral pressure coefficient are large. Wang et al. [8] quantitatively characterized the “space–time” process of energy release from microseismic events in rockbursts of different intensities and stated that microseismic activity showed “conduction and aggregation” along two intersecting structural planes and eventually formed clusters. Pan et al. [9] performed an in-depth analysis of the inherent response characteristics of the localized deformation and rockburst of mined coal-rock, further disclosing the crucial conditions for the occurrence of rockbursts and the response to the occurrence of localized deformation of mined coal-rock.
Regarding rockburst during the mining of steeply inclined coal seams, on the basis of the geodynamic zoning method, Zhang et al. [10–11] observed that the compressive strain generated by strong tectonic stress in the Wudong mining area is the energy source of rockburst and determined the critical energy density in two deep, steeply inclined working faces. He [12–13] found that the pressure stress of steeply inclined coal seams is the basic static load force source that induces rockburst. Lai et al. [14–16] reported that the coal body in the excavation-disturbed area of steeply inclined coal seams under strong earthquakes in the northwest mining area has clear localization characteristics. They believe that under the complex environment and multifield coupling of steeply inclined coal and rock masses, asymmetric deformation will take place, resulting in localized instability and further strong dynamic disasters. Based on this, the group [17] elucidated the mechanism of the “squeeze-prying” effect of steeply inclined coal-rock structures and further disclosed the impact of sharply inclined coal- rock through microseismic monitoring. As the mining depth increases, the seismic center of the rock pillar shifts from the middle to the two coal seams. Cui et al. [18] observed that the energy storage of steeply inclined coal rocks in the Wudong Coal Mine has a nonlinear positive correlation with mining depth, and the multivariate microseismic indicators before mine earthquakes all exhibited clear abnormal response relationships.
In summary, extant findings have undertaken a comprehensive analysis of the harmful consequences originating from coal-rock mining, particularly delineating the disaster-causing mechanisms inherent in rockbursts during the exploitation of steeply inclined coal-rock formations. However, many of these outcomes mainly use mechanical models to examine carefully the crucial role played by the rock pillar in disasters induced by rockbursts within sharply inclined coal seams. It should be mentioned that these studies neglected to elucidate the intricate “sandwiched rock pillar-roof” composite structure inherent to the mining of deep, steeply inclined coal seams and its contributory role in the disaster-evoking process. As the mining depth of steeply inclined coal seams has progressively increased recently, the related disaster induction mechanism and manifestation process have become more intricate. Both the sandwiched rock pillar and the roof of the working face show nonlinear mechanical response behaviors during rockbursts. Accordingly, our work is aimed at analyzing the mechanical response characteristics of pivotal disaster-prone regions and the spatiotemporal evolution features inherent in extensive microseismic data. By exploring deformation localization and shock response traits based on deep, steeply inclined coal-rock mining, we endeavor to unveil the mechanism that governs deformation localization and related thrust induction in steeply inclined coal-rock formations.
The Wudong Coal Mine is a prototypical mine in the Urumqi mining area and is characterized by steeply inclined and extremely thick coal seams. In the southern mining region, the average thicknesses of the B1+2 coal seam and the B3+6 coal seam are 35 and 45 m, with both 87°. The two coal seams exhibit weak bump proneness. A robust and exceptionally thick rock pillar is sandwiched between the two coal seams, boasting an average thickness of 100 m. The mining operation within the mine is organized based on the vertical orientation of the coal seam, with the working face length corresponding to the horizontal thickness of the coal seam. Sequentially, at the same horizontal level, the B3+6 coal seam is extracted before the B1+2 coal seam. After completing mining at each level, the mine uses backfilling materials such as loess and silt to fill the void created above the working face. Over extended periods of mining activity, this region has undergone the formation of a covering structure composed of a complex combination of loess, residual coal, fragmented overburden, and yellow sand [19–20].
The in-situ stress test results disclose strong horizontal tectonic forces exerted on the southern mining area. The prevailing stress field is characterized by a dominance of horizontal tectonic stress that measures 1.74–1.90 times the magnitude of vertical stress. Currently, the south mining area has been mined to the +425 level B3+6 working face, located approximately 375 m below the surface (Fig. 1). This working face has a mining height of 25 m, with a composition of the upper segment designated as the goaf and the lower section that is unmined solid coal at the +400 level. Notably, during mining, a presplitting process is used on the upper coal seam.
Seven severe shock accidents took place after mining to the +500 level (310 m depth) in the south mining area of the Wudong Coal Mine. This is much smaller than the average depth (600–800 m) at which the rockburst emerges in the western mining area. After positioning (Fig. 2), the source of the five accidents occurred in the sandwiched rock pillar. As the mining downward from the +500 level, the earthquake source position moves closer to the B3+6 coal seam with increasing mining depth. One case took place at the junction of the sandwiched rock pillar and the goaf; another occurred on the B6 roof. Analysis of the microseismic data in the 30 d before the seven shock accidents determined that the area with the highest frequency was the sandwiched rock pillar (43%), followed by B6 roof (27%), B3+6 coal seam (14%), and B1 floor (9%), and the frequency of microseismic events in the “sandwiched rock pillar-roof” area comprises a total of 70%.
The southern mining area of the Wudong Coal Mine uses the ARAMIS M/E system to perform real-time monitoring of microseismic events throughout the mining cycle. Analysis of microseismic events during 2013–2019 found that the sandwiched rock pillar attained the highest frequency of microseisms, followed by the B6 roof, B3+6 coal seam, B1 floor, and B1+2 coal seam. The sandwiched rock pillar is the area of most concentrated microseismic events. Events with microseismic energies greater than 105 J in this area comprise 47% of the total proportion of the stope and are the main release areas of the stope energy. With gradually increasing mining depth, high-energy microseismic events in the B6 roof area also gradually increase. Since 2013, 87% of microseismic events and 74% of high-energy events (energy greater than 105 J) have been distributed in the B3+6 coal seam and its roof and floor. Microseismic events greater than 107 J are also distributed in the B3+6 coal seam and its roof. Among them, the proportion of microseismic events greater than 106 J detected in the area between the rock pillar and the B6 roof accounted for 71.4%.
The above analysis suggests that the rockburst of the southern mining area shows clear zoning characteristics. The special coal-rock occurrence structure of the “sandwiched rock pillar-B6 Roof” is the source area of the rockburst. Analysis of the dynamic response characteristics of coal-rock in this area is urgently needed to implement targeted control measures on site.
With the extraction of sharply inclined coal seams, the sandwiched rock pillar and B6 roof have gradually become isolated and suspended. They experience bending deformation under horizontal tectonic stress. This state of existence makes it difficult for the sandwiched rock pillar and the B6 roof to collapse on a large scale. Thus, the curved, steeply dipping coal rocks squeeze the solid coal. The stress concentration zones in solid coal make it easier to reach the ultimate static load state. This gives high static stress conditions for the rockburst. Among them, the prying and squeezing caused by the rock pillar that is suspended on both sides will further increase the mining stress of the B3+6 coal seam, thus increasing the probability of rockburst in the coal-rock mass near its interface. The location of the earthquake source revealed by historical shocks also exhibits this point (the B3 roadway is the main disaster area). Thus, the prying and squeezing effect caused by the sandwiched rock pillar and the B6 roof is one of the main factors that cause rockbursts in the B3+6 coal seam.
The rockburst is the result of the superposition of a high static load state and a dynamic disturbance [21]. There are clear differences in the physical and mechanical properties of the coal-rock masses in the southern mining area. Compared with coal seams, the roof and floor are more suitable for energy storage and stress conduction. Thus, during the mining process of the B3+6 coal seam, the energy release resulting from the local rupture of the hard and extra thick sandwiched rock pillar and the bending and breakage of the roof will produce strong dynamic disturbances. The propagation of disturbance to the mining face will cause rapid changes in the stress of the coal mass. This can easily cause rockbursts. Thus, besides the high static stress state caused by the sandwiching of the rock pillar and the B6 roof, strong disturbance conditions are also one of the main factors causing the rockburst.
The rock pillar has greater rigidity and produces small bending deformations as mining activities continue. This leads to localized damage but is less likely to produce a global fracture. It is suggested that this small bending deformation is caused by the toppling and prying rotation of the rock pillar in the goaf. Only when the mining depth reaches a certain value will the prying, turning, and squeezing effect of the rock pillar on the coal seam be reflected. Thus, in this mechanical analysis, the tilting of the rock pillar is a plane strain problem, which is simplified into a beam structure mechanics model for derivation. Then, it is analyzed when the mining depth reaches where the rock pillar will show a strong squeezing and prying effect on the coal seam.
This model construction considers the middle of the rock pillar of the +450 B3+6 working face as the coordinate origin. The x-axis direction is the tendency of the rock pillar, and the y-axis direction is the vertical direction of the rock pillar. Stress analysis of the rock pillar during mining revealed that the rock pillar was supported to the right by the filling body above the B1+2 working face and the pressure from the top of the B3+6 working face during the goaf. The leftward supporting squeezing force of the filling body, the weight of the filling body, and the self-weight of the rock pillar occur. In the B3+6 working face area at the +450 level, the rock pillar is under rightward tectonic stress generated by the unmined B1+2 coal seam and the horizontal support force brought by the coal seam (which can be regarded as elastic foundation constraints). The B3+6 coal seam generates leftward tectonic stress, B3+6 coal seam gravity, and rock pillar gravity at this stage. The specific force diagram and simplified process of the rock pillar are shown in Fig. 3.
Because the rock pillar is in a discontinuous state of stress, we divided it into two stages for analysis. The slope length from the surface to the B3+6 working face is L, the slope length from the surface to the top of the B3+6 coal seam is a, and the stage slope length is m. Because the overall deformation trend of the rock pillar is to squeeze to the left in the first stage, the resultant force holding the rock pillar in the range of m ≤ x ≤ L is the support extrusion force of the filling body above the B3+6 working surface, the weight component of the filling body, and the difference between the gravity component of the rock pillar and the support extrusion force of the filling body above the B1+2 working surface on the left. In the second stage, the resultant force holding the rock pillar in the range of 0 ≤ x ≤ m is the tectonic stress transmitted by the B3+6 coal seam, the gravity component of the B3+6 coal seam, the gravity component of the rock pillar in this stage, the conduction of the B1+2 coal seam, and the difference between tectonic stress and its supporting extrusion force; the left side of the rock pillar can be regarded as an elastic foundation constraint state.
Combined with the above analysis, it can be found that the magnitude of the resultant force F0 at x = L and the magnitude of the resultant force F1 at x = m are
| F0=γRockPillarhcosθ | (1) |
| F1=γFillingacosθ+γRockPillarhcosθ | (2) |
The magnitude of F2 suffered by the rock pillar at x = 0 is
| F2=γCoalmcosθ+γRockPillarhcosθ | (3) |
Based on the above mechanical analysis and material mechanics content, it can be concluded that the distributed forces on the beam in the two stages are
| Q1(x)=F1−F0a(L−x)+F0 | (4) |
| Q2(x)=F1−F2mx+F2 | (5) |
where h is the thickness of the beam structure (100 m), θ is the inclination angle of the coal-rock (87°), γRockPillar is the bulk density of the rock pillar (24.83 kN/m3), and γFilling is the filling volume of the goaf area. The average bulk density of filling body is 19 kN/m3. γCoal is the average bulk density of the coal seam (13 kN/m3), and F0, F1, and F2 are the resultant stress of the rock pillar under load at different stages (Pa).
Comprehensive analysis of the stress state of the rock pillar in the range of m ≤ x ≤ L (stage one) is conducted by using the material mechanics section method. At this stage, the bending moment and deflection distribution of the cantilever beam at each position are
| M1(x)=16(2F0+F1)(L−x)2 | (6) |
| y1(x)=−172EJ(2F2+F1)(L−x)4+Ax+B | (7) |
where E is the elastic modulus of the beam (14 GPa), J is the moment of inertia (m4), and A and B are constants.
The rock pillar can be considered as an elastic foundation beam within the range of 0 ≤ x ≤ m (stage 2). Based on elastic foundation theory, the differential equation of the deflection curve corresponding to this section of the beam is
| d4y2(x)dx4+4β4y2(x)=Q2EJ | (8) |
where β = [k/(4EJ)]1/4 is a constant. Solving the above differential equation, the beam deflection equation is
| y2(x)=eβx[Ccos(βx)+Dsin(βx)]+e−βx[Ncos(βx)+Psin(βx)]+Q2k | (9) |
where C, D, N, and P are constants; because the constraint of the coal mass on the rock layer is an elastic foundation, k is the foundation coefficient (1000e6 kN/m3). Under the loading and constraint conditions of the mechanical model, when x approaches −∞, its deflection approaches constant, and e−βx approaches infinity. From this, we can obtain N = P = 0, and the bending moment of the stage-two elastic foundation beam is
| M2(x)=2EJβ2eβx[Csin(βx)−Dcos(βx)] | (10) |
In the x = m state, the beam meets the conditions of equal deflections, equal curvatures, and equal bending moments. The simultaneous solution draws the evolution process of the bending moment of the rock pillar, as shown in Fig. 4.
Combining our on-site measurements with the assignment of different parameters from another work [22], the calculated x0 (the extreme point of the bending moment) is 52.76 m, which is the +502.76 level. This is also the critical level at which significant extrusion and prying effects take place in the rock pillar. The force on the right side of the rock pillar under the influence of mining is significantly higher than that on the left side. As the mining activity continues, its tendency to bend to the left produces a bending moment. Because of the effect of the restraint boundary at the bottom end, the elastic foundation restraint generated by the coal mass reduces the bending of the second section beam to a certain degree. That is, the degree of prying and rotation of the second section of the rock beam is significantly lower than that of the first section of the beam. The maximum bending moment of the rock pillar is not at the fixed end of the origin or at the working surface. Under the constraints of the elastic foundation of the beam, there is an extreme bending moment at x = 52.7606 m in the second section of the beam, with a magnitude of 1038 kN·m. Small-scale rupture and instability are prone to take place. With increasing x, the force gradually decreases, and the bending moment also decreases, reaching 0 at the free end. This reveals that when mining reaches the +500 level (the +502.76 level is within the +500 B3+6 working face), the rock pillar is prone to mechanical instability, which in turn causes the coal-rock to be in a high static stress state.
The extreme bending moment point of the beam structure is the failure and instability position of the rock mass. The total elastic deformation energy of the entire beam is
| U(x)=∫M2(x)2EJdx | (11) |
The rate of change in strain energy with x is
| u(x)=U′(x)=M2(x)2EJ | (12) |
Fig. 5 displays the evolution trend of elastic energy in the rock pillar at different mining depths within the southern mining area. The origin of the coordinate system in this model is at the midpoint of the +450 level rock pillar. Notably, the energy growth rate undergoes a substantial surge in the range of +450 to +502 levels. Specifically, the elastic energy in the rock pillar peaks at 6.7 × 106 J at the +500 level, demonstrating a distinct deviation from the energy evolution pattern observed at Earth’s surface and extending to the +502 level. This divergence suggests that the energy accumulation during mining from the surface to the +500 level working face is relatively moderate, with the potential for effective energy release during the initial stages of rock pillar deformation localization. However, the energy accumulation dynamics below the +500 level working face, particularly at the +502 level, experience an abnormal acceleration due to the cumulative effects of tectonic stress, mining intensity, and coal and rock occurrence. In the deformation localization area, both strain growth and growth rate show significant increases, contributing to the observable delineation of the macroscopic failure interface development trend area.
From a mechanical standpoint, it is postulated that beyond the mining depth of the +502.76 level, the rock pillar experiences a dynamic levering phenomenon, with the high-energy zone between the rock pillar and the coal seam acting as the fulcrum. This scenario leads to the establishment of concentrated stress in the contact area between the rock pillar and the coal seams. Currently, the rock pillar encounters a transformative process, transitioning from the initial random distribution of local strain instability points in the shallower stages to a macroscopic catastrophic damage interface characterized by deformation accumulation. During this phase, the disturbance load arising from small-scale ruptures may induce a high-intensity rockburst.
The previous analysis revealed that the occurrence of rockbursts in the southern mining area was influenced by multiple factors, such as strong tectonic stress in the Urumqi mining area, i.e., the contact area between steeply inclined coal seams, rock pillar, and the roof will be in a state of highly concentrated stress due to the continuous “squeezing and prying” of the roof and rock pillar. The previous section presented an analysis of the prying dynamics behavior of the rock pillar at the critical mining depth. As a suspended mining roof, it will also produce more significant squeezing behavior on the coal seam after a certain mining depth, which comprehensively causes the “sandwiched rock pillar-B3+6 coal seam-roof” coal and rock system to be in a state of high static stress. Under the action of high-intensity horizontal segmented mining, small bending instability of the roof will generate a certain degree of energy release. After it is transmitted to the high static stress area of the coal and rock system as a dynamic load disturbance, it is easy to cause a rockburst in steeply inclined mining coal-rock under the superposition of high static stress and strong disturbance. Thus, as a key part of the force source in the southern mining area, the mechanical behavior of the B3+6 coal seam roof at different mining depths is also a crucial issue that needs an urgent solution.
Analysis of the direct jacking force during mining of the B3+6 working face revealed that the direct jacking in the goaf was supported and squeezed to the right by the filling body above the B3+6 working face. In the B3+6 working face, the B3+6 coal seam directly bears the tectonic stress to the right and the horizontal support force brought by the coal seam. The right side of the direct roof is influenced by the tectonic stress transmitted to the left by the deep roof, such as the main roof, gravity of the old roof, and gravity of the direct roof. The specific stress and simplified process of the coal seam roof in this case are illustrated in Fig. 6.
Because the overall deformation trend of the rock pillar is to squeeze to the left, the resultant force holding the rock pillar in the m ≤ x ≤ L range in the first stage is the support extrusion force of the filling body above the B3+6 working surface, weight component of the filling body, and the difference between the gravity component of the rock pillar and the support extrusion force of the filling body above the B1+2 working surface on the left. The resultant force of the rock pillar in the second stage in the m ≤ x ≤ L range is the tectonic stress transmitted by the B3+6 coal seam, gravity component of the B3+6 coal seam, gravity component of the rock pillar, conduction of the B1+2 coal seam, and the difference between tectonic stress and its supporting extrusion force.
Combined with the above analysis, the magnitude of the resultant force F0 at x = L and the magnitude of the resultant force F1 at x = m are
| F0=γimmlcosθ+γmainacosθ | (13) |
| F1=k2γmainacosθ+γimmlcosθ+γmainacosθ−λγsupacosθ | (14) |
The magnitude of F2 suffered by the rock pillar at x = 0 is
| F2=k2γmainLcosθ+γimmlcosθ+γmainacosθ−k1γimmLcosθ | (15) |
From the above mechanical analysis and material mechanics content, it can be concluded that the distributed forces on the beam in the two stages are
| Q1(x)=F1−F0a(L−x)+F0 | (16) |
| Q2(x)=F1−F2mx+F2 | (17) |
where l is the thickness of the roof (20.1 m), θ is the inclination angle of the coal-rock in the south mining area (87°), γimm is the average bulk density of the immediate roof of the B3+6 coal seam (26.67 kN/m3), γmain is the average bulk density of the main roof (29.82 kN/m3), and γsup is the average bulk density of the filling body in the goaf (19 kN/m3). The filling support coefficient λ changes with the depth of the roof. The support function at the surface is 0, and that at the working surface is 1; the overall application is linear. k2 is the tectonic force normally transmitted through the intact hard rock mass (1.8), k1 is the tectonic force transmitted through the fractured coal mass (1.5), and F0, F1, and F2 are the resultant stress of the roof at different stages (Pa).
The specific solution process for the bending moment of this mechanical model is similar to the solution process for sandwiching rock pillars and will not be repeated here (Fig. 7).
x0 is calculated to be 59.83 m, which is the level of +509.83, the critical level for the ultimate roof breakage. Due to the combined effect of tectonic stress and mining activities, the force on the right side of the direct top of the B3+6 coal seam is significantly higher than that on the left side. As mining activities continue, its tendency to bend to the left produces a bending moment.
Due to the effect of the restraint boundary at the bottom, the presence of the elastic foundation constraint generated by the coal mass reduces the bending of the second section beam to a certain degree, that is, the degree of prying and rotation of the second section beam directly supported is significantly smaller than that of the first section beam. Therefore, the bending moment extreme value of the direct top overall is that there is an extreme bending moment value at x = 59.83 m in the second section of the beam under the constraints of the elastic foundation of the beam. The magnitude is 993 kN·m, which is prone to breakage. With increasing x, the force gradually decreases, and the bending moment also decreases, reaching 0 at the free end. The evolution trend of the roof shear force also reflects this phenomenon. This shows that when mining reaches the +509 level (the +509.76 level is within the +500 B3+6 working face), the B3+6 coal seam is prone to large fracture instability when directly jacked, causing dynamic load disturbance.
The evolution trend of the direct roofing elastic energy of the B3+6 coal seam at various mining depths in the southern mining area is plotted in Fig. 8. The origin of the coordinate system of this model is taken from the middle of the direct top at the +450 level, and the energy growth rate of the direct top increases sharply from the +450 to +509 level. At the +509 level, the rock pillar energy can reach 3.1 × 106 J, which is the same as the +509 level; however, the energy evolution situation on the surface is completely different. That is, the degree of energy accumulation is not large when mining from the surface to the +500 horizontal working face, and the accumulated energy may be effectively released during the localization process of the roof deformation in the initial stage. However, due to the multiple influences of tectonic stress, mining intensity, and coal and rock occurrence, the roof below the +500 horizontal working face (+509 level) has abnormally accelerated energy accumulation, and the strain growth and growth rate are obvious in the localized deformation area. This increase results in the gradual clarity of the development trend area of its macroscopic damage interface.
From a mechanical perspective, we believe that when the mining depth is higher than the +509.83 level, the coal seam roof will experience an obvious overall fracture, and the strong extrusion effect on the coal seam will cause the contact area between the roof and the coal seam to be in a state of extremely high concentrated stress. At this time, the disturbance load generated by the fracture is sufficient to induce a high-intensity shock.
In conclusion, the sharply inclined mining of coal-rock in the Wudong Coal Mine shows obvious deformation localization with increasing mining depth. The +500 level is the deformation localization inflection point. After the +500 level, the rock pillar and the B6 roof are prone to fracture and instability to varying degrees. The contact area between the B3+6 coal seam, the rock pillar, and the roof has been in a state of highly concentrated stress for a long time and is easily prone to rockburst due to dynamic disturbances.
Domestic and foreign scholars have found the phenomenon of deformation localization in disaster manifestations of different scales (e.g., indoor sample rupture, coal burst, and tunnel rockburst) and believed that its positive correlation and strong traceability can be employed to further study its relationship with engineering. The response relationship of geological hazards [23] offers a theoretical basis for predicting engineering geological hazards based on deformation localization. According to microseismic monitoring and analysis of long-term observation results of large-scale geological disasters such as rockbursts and landslides, it is believed that there is a significant abnormal accumulation of on-site microseismic events before the appearance of such geological disasters [24]. Meanwhile, it is suggested that the prediction effect of localized deformation on rockburst can be further enhanced based on relevant data such as microseismic data.
According to the first law of geography, Professor Patrick of Oxford University put forward a quantitative study of the degree of spatial data aggregation by using spatial autocorrelation measurement, which has been gradually applied in the field of engineering disaster analysis [25]. Thus, we attempt to integrate the relevant indicators of the spatial autocorrelation theory into the analysis of dynamic damage response patterns of mined coal and rock and then use massive microseismic data to examine the degree of localization deformation of steeply inclined mining coal-rock to generate a more comprehensive report on the “spatiotemporal” response characteristics of dynamic damage of coal and rock resulting from sharply inclined mining in strong earthquake areas through the microseismic “clustering coefficient” under the disturbance response in the southern mining area.
Spatial autocorrelation statistics is a basic tool to measure responses between geophysical data that may no longer be independent but correlated because they are influenced by spatial diffusion interactions. Thus, spatial autocorrelation is based on statistical methods by investigating certain response characteristic values between a certain spatial unit and its surrounding units in a spatial area to analyze the characteristics of the spatial distribution phenomenon of these spatial units. Currently, the most commonly used global analysis index in spatial statistics is the global Moran index (Moran’s I index). If it is determined that this phenomenon has aggregation characteristics on a spatial scale, you can choose to use local correlation metrics to reflect the impact of this attribute variable on spatial agglomeration in the local area. The most used local analysis index is the hotspot analysis index (Getis-Ord Gi* index). The global Moran’s I calculation method is
| Moran’sI=1∑ni=1∑nj=1Wij×n∑ni=1∑nj=1Wij(xi−ˉx)(xj−ˉx)∑ni=1(xi−ˉx) | (18) |
where n is the number of regional units, xi is the attribute observation value of regional unit i, and Wij is the spatial weight coefficient. When areas i and j are adjacent, Wij = 1; if not, Wij = 0.
The value range of Moran’s I coefficient is [−1, 1]. When its value is between 0 and 1, the spatial data shows a positive correlation, and the larger the value, the stronger the spatial agglomeration. When it is between −1 and 0, it shows a negative correlation, and the smaller the value, the stronger the spatial discreteness. A value of zero is when it appears as a random distribution.
The expected value of Moran’s I is
| En(I)=−1n−1 | (19) |
The variance of Moran’s I VARn(I) is
| VARn(I)=n[(n2−3n+3)s1−ns2+3s20]−λ[(n2−n)s1−2ns2+6s20]s20(n−1)(n−2)(n−3) | (20) |
Among them, s0=∑ni=1∑nj=1Wij; s1=12∑ni=1∑nj=1(Wij+Wji)2; s2=∑ni=1(Wi+Wj)2; Wi and Wj is the sum of the ith and jth columns.
For Moran’s I, the significance can be calculated using the statistic Z:
| Z=I−En(i)VARn(i) | (21) |
We verify whether to accept the hypothesis based on the 5% confidence interval. When the spatial data is clustered, Z > 1.96; when the spatial data is distributed, Z < −1.96. As an optimization index for high/low clustering (Getis-Ord General G), the hot spot analysis index is calculated using the binary weight method and can invert the local spatial agglomeration density peak.
| G∗i=n∑nj=1Wijxj−∑nj=1xj∑nj=1Wijn√∑nj=1x2jn−(∑nj=1xjn)2 | (22) |
Based on the microseismic event data in the southern mining area from August 2015 to October 2016, the spatial autocorrelation index of the microseismic data at this stage was calculated and is shown in Table 1. The results reveal that the microseismic data in the southern mining area have obvious agglomeration; thus, it is urgent to investigate whether the inclined mining coal rocks show significant localized fractures in the strike direction.
| Microseismic frequency |
Stage average Moran’s I |
En(I) | VARn(I) | Z |
| 10669 | 0.5183 | −0.0205 | 0.0064 | 7.6351 |
DownLoad:
CSV
By calculations, it can be found that the microseismic data under the influence of mining in the southern mining area has obvious aggregation on the spatial scale. Thus, on the basis of the microseismic data at this stage, the time when the microseismic event on the day is zero due to production factors is removed. We conduct an in-depth analysis of multiple information such as the b value of the remaining effective time, large energy events of the day, cumulative microseismic frequency, daily microseismic energy, and microseismic local autocorrelation indicators. Afterward, we obtain the coal and rock dynamic parameters and local space of sharply inclined mining in the Wudong Coal Mine. The autocorrelation response relationship is illustrated in Fig. 9. The red and purple circles in the upper and lower figures, respectively, reflect the different responses of the “space–time” characteristics in this stage, and their ranges refer to the slope distortion of the cumulative microseismic frequency curve. When many large energy events (107 J) occur during this stage, the frequency of microseisms on that day is low and is positively correlated with the b value and negatively correlated with the spatial agglomeration coefficient. The agglomeration coefficient values are all between 0.6 and 1. Meanwhile, the b value shows a downward trend before the high-energy event occurs and attains its lowest value in recent days on the day of the high-energy event. However, the spatial agglomeration coefficient is opposite to the evolution trend. It slowly increases before the high-energy event occurs and reaches a peak value in recent days on the day of the high-energy event. The agglomeration coefficient always exhibits a high value with a small amplitude near the b value shock range. The low b value reflects the highly concentrated stress state of the rock at this stage and is a typical precursor indicator of rock fracture. On an engineering scale, this means that the disturbed rock mass is in a critical period for strong earthquakes. At this time, the spatial aggregation coefficient is also at a stable high value, and it can be regarded that the deformation localization phenomenon gradually takes place in the engineering disturbed rock mass. Meanwhile, we found that as mining activities continue, although the spatial aggregation coefficient still responds to the b value evolution law, its average value slowly increases, demonstrating a positive correlation between the spatial agglomeration of microseismic events and mining activities. The six black crosses in the figure indicate the cumulative frequencies of microseisms, all of which emerge close to the area where high-energy events induce a decrease in the b value. We found that after the 107 J energy level event, the b value has a relatively obvious rebound phenomenon, which is also similar to the evolution characteristics of the b value near the day of a large magnitude earthquake in seismology.
Thus, we believe that the mined coal-rock has undergone localized deformation and damage under high-intensity mining. As mining activities continue, the average clustering coefficient gradually increases, showing that the local damage situation is gradually intensifying, and there may be one or several coal rocks in this region in an accelerating state of “expansion–penetration” of fissures, and the probability of macroscale fissure penetration and instability increases significantly. Facts also show that a 9.5 × 106 J shock accident did occur in November 2016, which further confirms the macroscopic disaster-causing impact of the localized deformation of mining coal and rock. There is a certain positive correlation between the spatial agglomeration of microseismic events and the localized damage caused by the mining of coal and rock. Before the occurrence of high-energy events, the b value gradually decreases while the spatial agglomeration coefficient increases, and there is a certain negative response relationship between the two. From the above conclusions, if the response relationship between the abovementioned b value, large energy event, and spatial agglomeration coefficient in steeply inclined coal seams is detected in real time, it is feasible to further determine the rockburst risk of steeply inclined coal seams.
The principle of rockbursts induced by the superposition of dynamic and static loads has gained widespread interest both domestically and internationally [26]. The static stress field of the surrounding rock experiences energy “transmission–release” following localized damage to the coal-rock. When the superimposed dynamic load surpasses its critical shock load, a rockburst occurs (Fig. 10). This prevailing condition reduces the likelihood of extensive collapse in both the roof and rock pillar. During the relatively shallow phases, the need for backfilling leads to limited void space. However, the fill material is enough to offer substantial support to the rock pillar and roof. As a consequence, the structural stresses at this stage are insufficient to induce roof fracture. Internal strain growth within the coal-rock is characterized by a relatively small and randomly distributed increase, enabling effective energy release as microscopic fissures develop. As mining activities continue to greater depths, localized deformation in the coal-rock speeds up. Simultaneously, the fill material becomes inadequate to satisfy the practical demands of the void space. The uneven collapse and accumulation of gangue in the goaf lead to nonuniform loading on the underlying mining space. The increasing structural stresses and decreasing support constraints of the fill material collectively contribute to a dual influence of “squeezing–levering” on the B3+6 coal seam beneath the rock pillar and roof, causing a pronounced high concentration of stress. During this phase, the dynamic disturbances induced by the limited-range roof fractures easily result in rockbursts.
With the continuous development of deep, steeply inclined working faces, the upper loose overburden in the goaf descends continuously to fill the void, leading to a continual increase in the suspended area of the roof and floor. The rock pillar shows a dynamic levering condition pivoted on the high-energy zone where it contacts the coal seam. This exacerbates the prying effect of the roof in the strike direction, with the stress concentration position serving as the “fulcrum.” In the dip direction, significant differences in the mechanical parameters of the loose overburden accumulated behind the goaf and the unmined solid coal in front contribute to noticeable variations in their support functions on the surrounding rock. The weaker supporting capacity of the loose overburden makes it demanding to establish effective support under mining-induced stress, giving rise to a sudden increase in the deformation of the surrounding rock. Meanwhile, the unmined coal shows a relatively robust supporting capacity, causing a smaller deformation in the surrounding rock. The incongruent deformations of the surrounding rock along the strike direction culminate in front of the working face, acting as a “fulcrum,” inducing prying effects in the strike direction.
Upon reaching the +500 level in the extraction process, the spatial superposition of the prying effects in the strike and dip directions intensifies. This worsens the energy distribution disparity and stress asymmetry of the “rock pillar-B3+6 coal seam-B6 roof,” a unique structure characterized by a “strong–weak” configuration. As a result, localized rock masses undergo tensional and shear failures. The fracture locations within the rock mass serve as high-energy dynamic sources that propagate stress waves to the surrounding rock. Together with elevated dynamic loads, this interaction with the extraction-induced effects in the goaf forms multiple disturbances. If these disturbances, compounded with the high static stress in the roadway, exceed the critical impact threshold, a rockburst will inevitably occur. Simultaneously, within the strike direction of the mining space, a substantial suspended roof remains unstable. The sandwiched rock pillar and the B6 roof, as the mining depth increases, are prone to accumulating energy. The B6 roof and rock pillar continue to exert pressure on the coal, and upon reaching a certain hanging length, fractures and slips happen. This process may cause the frequent occurrence of rockbursts. Hence, the fundamental cause of the frequent rockburst incidents in the B3+6 coal seam working face and recovery roadway lies in the combined response of the B6 roof fracture, localized rock pillar damage-induced dynamic disturbances, and high static stress in the coal seam, manifesting in the “strike-dip” direction.
The energy field distribution, dissimilation path, and conduction efficiency of steeply inclined mining are determined by the coupling effect of the physical and mechanical properties of the coal-rock medium and its tight integrity. In terms of physical and mechanical parameters, the rock pillar, B3+6 coal seam, and B6 roof construct a similar “strong–weak–strong” structure. The physical and mechanical parameters of the rock pillar, B6 roof, and coal seam are different. This leads to its energy storage capacity being larger than that of coal seams, and the energy and stress conduction efficiency of relatively complete and dense rock pillars and B6 roofs is also high. Thus, from the perspective of reducing the energy storage capacity of key rock mass and cutting off stress conduction paths, the “internal–external” collaborative modification control technology is divided into two aspects. “Inside” indicates using a combination of presplitting blasting and water injection for the central part of the rock pillar that plays a leading role: water injection is employed to decrease the energy storage capacity of the rock pillar area; blasting is utilized to release its accumulated elastic energy, thereby alleviating the rock. The strong extrusion force formed by the column exerts a high static load stress on the mining space. “Outside” indicates coordinated blasting of deep and shallow holes on the roof and floor of the B3+6 coal seam. Deep hole blasting can decrease the influence of the maximum horizontal principal stress and cut off the stress transmission path; shallow hole blasting can prevent large-scale slippage and collapse of shallow rock mass.
In the early stage, we obtained scientific and reasonable control and optimization parameters by using relevant three-dimensional fine numerical simulations [27], which will not be described again here. This section is based on the previous energy dissimilation path results and is combined with the actual situation on site to derive the corresponding energy control plan: (1) water injection blasting of rock pillar and stone gates: 1 and 2 chambers are constructed at 1650 m east and 1350 m west of the +425 level, respectively. Each chamber tends to have 2 rows of blasting holes arranged in a fan shape, 10 holes/row, a hole diameter of 113 mm, a hole length of 39–79 m, and a hole sealing length of 15–20 m. The construction length of the water injection hole is 135 m, the construction angle is 6°, the hole diameter is 113 mm, and the hole sealing length is 20 m. The holes are sealed with a hole sealer and reinforcement. (2) advanced deep and shallow hole blasting in the roof and floor of the B3+6 coal seam: The same control measures are conducted for the floor. Deep and shallow holes are arranged alternately, and a group of blasting holes is formed every 10 m, which are 2 holes/group and 3 holes, respectively. The row spacing of shallow blasting holes is 10 m/row, with 3 boreholes arranged in each group; the row spacing of deep hole blasting is 30 m/row, and 2 boreholes are arranged in each group. The specific parameters of the blasting holes are summarized in Fig. 11.
The transient electromagnetic (TEM) method is also called the time-domain electromagnetic method. The apparent resistivity of fractured coal-rock is higher than that of intact coal-rock [28–29]. In this work, TEM monitoring technology is employed on the +425 B3+6 working face to characterize the degree of coal-rock fracture based on TEM data. The monitoring principle and equipment are illustrated in Fig. 12.
TEM equipment was employed to detect the +425 leval rock pillar and the B6 roadway roof. Three measurement lines (0°, 30°, and 60°) were designed for the two tunnel detection processes, and the vertical direction of the tunnel was defined as 0 Spend. The detection position of the rock pillar is 1090–1120 m, and the detection position of the B6 tunnel roof is 1595–1540 m. The detection results of the rock pillar are presented in Fig. 13. The overall relative obvious resistivity of the unregulated area is small, demonstrating that the integrity of the rock surrounding this area is relatively good. The apparent resistivity range of the area where energy control measures are implemented is 30–105 Ω·m, and the average apparent resistivity is 430% higher than that before no control.
The B6 roof detection results are displayed in Fig. 14. The overall relative obvious resistivity of the unregulated area is small, with that of most areas being 25–30 Ω·m, showing that the integrity of the surrounding rock in this area is relatively good. The apparent resistivity range of the area where energy control measures are implemented is 75–100 Ω·m, and the average apparent resistivity is 300% higher than that before no control. An annular high-resistance zone is generated around the high apparent resistance area, showing that the control measures have maximized the loosening and cracking of the rock mass. The formed high-resistance damage zone of the surrounding rock shows that energy control has caused obvious damage to the rock surrounding the steeply inclined slope. The destruction effect is extremely obvious.
In this work, we investigated the mechanical response behavior of critical disaster-prone regions in sharply inclined working faces. We elucidated the interlinked mechanisms that govern the localized deformation of coal-rock in the sharply inclined extraction of the Wudong Coal Mine. Leveraging extensive in-situ microseismic data, we disentangle the “time–space” response relationship of the dynamic failure of coal-rock. The important conclusions can be drawn.
(1) The distinctive “sandwiched rock pillar-B6 roof” structure acts as the source region for rockburst in the southern mining area. The frequent occurrence of rockbursts during the extraction of steeply inclined extra-thick coal seams is ascribed to the collaborative effect of high static stress produced by the squeezing and prying action of the sandwiched rock pillar and B6 roof on the coal seam, together with the strong disturbance resulting from the localized rupture of the rock pillar and the fracture of the roof.
(2) The critical depth for the conspicuous squeezing and levering effect of the rock pillar was determined at the +502 level, while that for the ultimate rupture of the roof was at the +509 level. Thus, the +500 level serves as the critical mining depth for rockburst in the southern mining area. With increasing mining depth, both bending deformation and energy accumulation of the rock pillar and roof show nonlinear and rapid escalation. The localized deformation of deep, steeply inclined coal-rock leads to the spatial superposition of squeezing and prying effects, exacerbating the energy distribution difference and stress asymmetry of the “sandwiched rock pillar-B3+6 coal seam-B6 roof.” This significantly increases the threat of frequent rockbursts.
(3) Building on our developed high-energy distortion zone “inner–outer” coordinated modification and control technology, targeted control measures were applied for the identified key areas. TEM monitoring data from the field demonstrate that this control technology effectively reduces the high stress concentration in the rock surrounding the mining field and the energy distortion. As a consequence, it ensures the safe and efficient development of steeply inclined coal seams.
This work was financially supported by the Major Program of the National Natural Science Foundation of China (No. 52394191), the Outstanding Ph.D Dissertation Cultivating Program of Xi’an University of Science and Technology (No. PY22001), and the National Foundation for studying abroad (No. [2022] 87).
The authors declare that they have no conflicts of interest.
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| Microseismic frequency |
Stage average Moran’s I |
En(I) | VARn(I) | Z |
| 10669 | 0.5183 | −0.0205 | 0.0064 | 7.6351 |
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