# Mechanism of Coal Burst and Prevention Practice in Deep Asymmetric Isolated Coal Pillar: A Case Study from YaoQiao Coal Mine

Authors: Chengchun Xue; Anye Cao; Wenhao Guo; Songwei Wang; Yaoqi Liu; Zhiping Shen
Journal: Shock and Vibration
 (2021)

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## Abstract

Coal pillar bursts continue to be a severe dynamic hazard. Understanding its mechanism is of paramount importance and crucial in preventing and controlling its occurrence. The extreme roadway deformations from the asymmetric isolated coal pillars in the central mining district of YaoQiao Coal Mine have responded with frequent intense tremors, with risky isolated coal pillar bursts. The theoretical analysis, numerical simulation, and field measurements were done to research the impact of spatial overburden structure and stress distribution characteristics on the isolated coal pillar area, aiming to reveal the mechanism of coal pillar burst leading to the practice of prevention and control in the asymmetric isolated coal pillar area. The study shows that the overburden structure of the asymmetric is an asymmetric “T” structure in the strike-profile, and the stress in the coal pillar is mostly asymmetric “saddle-shaped” distribution, with the peak stress in the east side of the coal pillar, and the coal pillar is a “high stress serrated isolated coal pillar.” Numerical simulation results showed that the support pressure in the isolated coal pillar area on the strike profile was asymmetrically “saddle-shaped” distribution. The peak vertical stress in the coal pillar area continued to rise and gradually shifted to the mining district's deep part. As a result, the response of the roadway sides to the dynamic load disturbance was more pronounced. They developed a coal burst prevention and control program of deep-hole blasting in the roof of asymmetrical isolated coal pillar roof and unloading pressure from coal seam borehole. Monitored data confirmed that the stress concentration was influential in the roadway’s surrounding rock in the asymmetric isolated coal pillar area, circumventing coal pillar burst accidents. The research outcomes reference the prevention and control of coal bursts at isolated working faces of coal pillars under similar conditions.


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## Body

## 1. Introduction

Coal bursts, a dynamic geological hazard in underground coal mining, are characterized by coal and rock masses’ sudden and violent failure with a significant energy release. They posed a severe threat to the production and safety of underground coal mining or surface buildings and, in some cases, resulted in injury and loss of life [1, 2]. In recent years, with the increase in the depth and intensity of coal mining in China, mining conditions have deteriorated dramatically, and the frequency of coal burst accidents has a rising trend [3]. The occurrence of coal burst was closely related to the degree of stress concentration in the coal and rock layer and the overburden structure in the working face [4, 5]. With the gradual hollowing out of the working face on both sides of the mining district roadway, the protection coal pillar area gradually formed an isolated structure. With hollowing on both sides, the static stress load concentration of the roadway surrounding the rock system in the isolated coal pillar area is high. When it was affected by thick and hard rock layer breaking and working face mining disturbance, the isolated coal pillar area was affected by the superposition of concentrated load. The higher was the possibility of inducing coal pillar bursts [6]. For example, On March 27, 2014, a coal burst occurred during the return roadway excavation period in the 21032 mining district: six workers died and thirteen workers were injured [7]. On August 15, 2014, a coal burst occurred in 35000 mining district, where a 300 m long mining road was damaged, two workers were killed [8]. It was common in mine production to form isolated protection coal pillars after extensive hollowing on both sides of the mining district. Therefore, it was essential to study the mechanism of coal bursts in isolated protection coal pillar areas to protect the pedestrian and ventilation and transportation tasks in the mining district roadway.Chen et al. [9] studied the overlying rock movement and stress distribution over the equivalent mining height method over isolated coal pillars. Jia et al. [10] analyzed the overburden damage pattern and coal pillar deformation characteristics during mining the isolated working face under thick and hard critical seam conditions based on similar material simulation tests. Hao et al. [11] theoretically analyzed the rock movement characteristics and stress distribution pattern of the isolated working face, and the theoretical criterion for safe mining of the isolated working face was proposed. Fuxing et al. [12] analyzed the dynamic change process of the roof structure in irregular mining areas on both sides during work face retrieval. They presented the calculation formula of coal body support pressure and destabilization analysis method for an isolated working face. Li et al. [13] studied dynamic stress evolution in the isolated protective pillar during the long wall panel analysis using PFC2D. The computational model of abutment pressure was established by theoretically analyzing the stress transfer mechanism of the two types of roof structure, and the calculation method for abutment stress was proposed [14]. Reed et al. [15] studied the mechanical characteristics of the coal pillar system and proposed the interaction mechanism between coal pillars and overburden. Hamid [16] studied that, through the methods of field observation, geological monitoring, and numerical simulation, the horizontal stress accumulation in the coal pillar-topside-bottom plate was quantitatively analyzed. It was concluded that the horizontal stress in the high-stress coal pillar and the surrounding rock would cause impact and trigger the violent destabilization of the coal pillar.Many scholars researched the stability of isolated coal pillars from the overburden structure and stress distribution characteristics of the isolated working face (or isolated coal pillar) and made many research findings. Still, the research on coal burst induced by the instability of asymmetric isolated coal pillars (size asymmetry of coal pillar and size asymmetry of goaf on both sides) is rare. Moreover, coal burst hazards are subject to different coal pillars and various impact hazards, which are difficult to prevent and control. Thus, the coal burst accident induced by the overall instability of asymmetric isolated coal pillars is an emotional disaster problem that needs addressing in coal mine safety production.YaoQiao Coal Mine’s central mining district protecting coal pillar, on both sides of an extensive range of mining void and size irregularities, formed an asymmetric isolated coal pillar area. On August 25, 2016, a coal burst event (Figure1, red star area), magnitude up to 1.9 on the Richter scale, resulting in the return roadway over a 60 m range, of the return roadway, track roadway nearly 70 m range of the bottom of the drum, flake gang, roof cracking, the roadway damage as severe. Therefore, the paper analyzed the spatial overburden structure and stress distribution characteristics of the asymmetric coal pillar area in the mining district, explored the mechanism of coal burst occurrence in the asymmetric isolated coal pillar area, and formulated a coal burst prevention plan for the coal pillar area, aiming to provide a valuable reference for coal burst prevention in the isolated protected coal pillar area under similar conditions.Figure 1
Central mining district district’s isolated asymmetric isolated coal pillar arrangement: (a) actual measurement layout; (b) sketch diagram.
(a)(b)


## 2. Site Characteristics

### 2.1. Asymmetric Isolated Coal Pillar Details

For the asymmetric isolated coal pillar in the YaoQiao Coal Mine central mining district, the selected research project was in Xuzhou City, northwest Jiangsu Province, China. Based on preliminary statistics, at least two disastrous coal bursts had previously occurred in this YaoQiao Coal Mine, which has a high coal burst risk and is a characteristic coal mine with coal burst hazards.The central mining district of YaoQiao Coal Mine is endowed with No. 7 coal seam, having a simple endowment structure. The No. 7 coal seam thickness is 4.59 to 6.19 m (5.47 m on average) with an inclination ranging from 4° to 10° (8° on average); the coal seam has a substantial coal burst impact tendency and a significant mining depth of 650 to 850 m in the central mining district, and the original stress of coal and rock mass is high.The central mining district adopted double-wing mining, with gradual mining of the working face on both sides of the coal pillar in the main roadway; the district dip protective coal pillar in the mining area had formed an isolated asymmetric structure. The narrowest area of the isolated coal pillar was about 110 m, and the most expansive area was about 180 m. Meanwhile, a fault protection coal pillar of about 90 m in width and 450 m in length was left in the west wing goaf. Figure1 illustrates the mining district layout.We first determined the critical layer’s bearing level with the thick and hard roofed structure group having different thicknesses and the bearing of the multilayer roof composite structure meeting the relationship in Equation (1) [17]:(1)qnm=Emhm3∑i=mnγihi∑i=mnγihi3,where (qn)m is the load of the roof nth on the roof m; mth, n, and i are the roof rock sequence numbers, respectively; Em and hm are the elastic modulus and thickness of the mth overburden, respectively; and Ei, hi, and γi are the elastic modulus, thickness, and unit weight of the ith overburden, respectively.Thenth overburden layer satisfies the relationship in Equation (2) for the critical stratum [17]:(2)qnm<qn−1m.The log of central mining district borehole Y13# combined with the roof fracture data observed during the mining period of multiple working faces in the mining district helped determine that the sandstone group 17.0 m away from the coal seam and 26.34 m thick was inferior key strata 1, and the siltstone group 417.0 m away from the coal seam, and 71.34 m thick was the leading fundamental stratum. Meanwhile, the main roof and floor were classed with weak coal burst tendency through laboratory testing for coal burst tendency characterization. Figure2 gives the critical strata information.Figure 2
Distribution of main primary and inferior key strata in the asymmetric coal pillar area (borehole Y13#).


### 2.2. Analysis of Actual Field Measurement Data

Influenced by the wide range of goaf on both sides of the asymmetric isolated coal pillar, the stress concentration degree of the surrounding rock system in the mining area increased. As a result, the return roadway was affected by high static load stresses, and the shotcrete in most areas of the sides and the upper part of the roadway appeared cracked and hung. Moreover, the bottom drum phenomena occurred in the local area of the roadway floor, as shown in Figure3.Figure 3
Actual photo of the return roadway: (a) schematic diagram of sidewalls of the roadway sideways and floor heave; (b) schematic diagram of roadway roof cracking.
(a)(b)Microseismic monitoring in mines gave seismic event location and calculation of seismic energy to evaluate the mining stress state and coal burst hazard [18]. A microseismic monitoring system called SOS, developed by the Central Mining Institute of Poland, was installed in the YaoQiao Coal Mine. The maximum locating errors were 20 m in the horizontal direction and 30 m in the vertical direction. The seismic monitoring network from July to October 2018 is shown in Figure 1.The microseismic activity in the isolated coal pillar area was frequent due to design working face roadway excavation, which indicated that the coal and rock system in the coal pillar area was disturbed by mining, which intensified its stress evolution and rupture activity. Figure4 illustrates the distribution of microseismic events. Figure 4(a) shows that the microseismic event count and microseismic intensity in the west wing goaf of the central mining district were significantly more than those in east wing goaf; especially the microseismic activity was the most intense between the designed working face and the faulted coal pillar, indicating that the stability of the coal and rock layer was worse in this area. Furthermore, Figure 4(b) shows that microseismic events in the east wing goaf occurred mainly in the high rock layer, while those in the west wing goaf occurred primarily in the low rock layer, indicating that the overburden fissures in the east and west wing goaf of the asymmetric isolated coal pillar area were at different heights.Figure 4
Microseismic localization map of the asymmetric isolated coal pillar area (2018.01.01∼2018.06.30): (a) microseismic positioning plan; (b) microseismic positioning orientation profile.
(a)(b)With a precise picture of the roadway response in the coal pillar area and microseismic monitoring, the asymmetric isolated coal pillar in the mining area was affected by the coal seam mining on both sides. Furthermore, the stress concentration of the roadway surrounding the rock system was high, the overlying rock was highly active, and the roof's stability was unsatisfactory. Therefore, to reduce the safety of the roadway group in the isolated coal pillar area during the design working face excavation and later working face retrieval, it was imperative to study the mechanism of coal burst in the asymmetric isolated coal pillar area.


## 2.1. Asymmetric Isolated Coal Pillar Details

For the asymmetric isolated coal pillar in the YaoQiao Coal Mine central mining district, the selected research project was in Xuzhou City, northwest Jiangsu Province, China. Based on preliminary statistics, at least two disastrous coal bursts had previously occurred in this YaoQiao Coal Mine, which has a high coal burst risk and is a characteristic coal mine with coal burst hazards.The central mining district of YaoQiao Coal Mine is endowed with No. 7 coal seam, having a simple endowment structure. The No. 7 coal seam thickness is 4.59 to 6.19 m (5.47 m on average) with an inclination ranging from 4° to 10° (8° on average); the coal seam has a substantial coal burst impact tendency and a significant mining depth of 650 to 850 m in the central mining district, and the original stress of coal and rock mass is high.The central mining district adopted double-wing mining, with gradual mining of the working face on both sides of the coal pillar in the main roadway; the district dip protective coal pillar in the mining area had formed an isolated asymmetric structure. The narrowest area of the isolated coal pillar was about 110 m, and the most expansive area was about 180 m. Meanwhile, a fault protection coal pillar of about 90 m in width and 450 m in length was left in the west wing goaf. Figure1 illustrates the mining district layout.We first determined the critical layer’s bearing level with the thick and hard roofed structure group having different thicknesses and the bearing of the multilayer roof composite structure meeting the relationship in Equation (1) [17]:(1)qnm=Emhm3∑i=mnγihi∑i=mnγihi3,where (qn)m is the load of the roof nth on the roof m; mth, n, and i are the roof rock sequence numbers, respectively; Em and hm are the elastic modulus and thickness of the mth overburden, respectively; and Ei, hi, and γi are the elastic modulus, thickness, and unit weight of the ith overburden, respectively.Thenth overburden layer satisfies the relationship in Equation (2) for the critical stratum [17]:(2)qnm<qn−1m.The log of central mining district borehole Y13# combined with the roof fracture data observed during the mining period of multiple working faces in the mining district helped determine that the sandstone group 17.0 m away from the coal seam and 26.34 m thick was inferior key strata 1, and the siltstone group 417.0 m away from the coal seam, and 71.34 m thick was the leading fundamental stratum. Meanwhile, the main roof and floor were classed with weak coal burst tendency through laboratory testing for coal burst tendency characterization. Figure2 gives the critical strata information.Figure 2
Distribution of main primary and inferior key strata in the asymmetric coal pillar area (borehole Y13#).


## 2.2. Analysis of Actual Field Measurement Data

Influenced by the wide range of goaf on both sides of the asymmetric isolated coal pillar, the stress concentration degree of the surrounding rock system in the mining area increased. As a result, the return roadway was affected by high static load stresses, and the shotcrete in most areas of the sides and the upper part of the roadway appeared cracked and hung. Moreover, the bottom drum phenomena occurred in the local area of the roadway floor, as shown in Figure3.Figure 3
Actual photo of the return roadway: (a) schematic diagram of sidewalls of the roadway sideways and floor heave; (b) schematic diagram of roadway roof cracking.
(a)(b)Microseismic monitoring in mines gave seismic event location and calculation of seismic energy to evaluate the mining stress state and coal burst hazard [18]. A microseismic monitoring system called SOS, developed by the Central Mining Institute of Poland, was installed in the YaoQiao Coal Mine. The maximum locating errors were 20 m in the horizontal direction and 30 m in the vertical direction. The seismic monitoring network from July to October 2018 is shown in Figure 1.The microseismic activity in the isolated coal pillar area was frequent due to design working face roadway excavation, which indicated that the coal and rock system in the coal pillar area was disturbed by mining, which intensified its stress evolution and rupture activity. Figure4 illustrates the distribution of microseismic events. Figure 4(a) shows that the microseismic event count and microseismic intensity in the west wing goaf of the central mining district were significantly more than those in east wing goaf; especially the microseismic activity was the most intense between the designed working face and the faulted coal pillar, indicating that the stability of the coal and rock layer was worse in this area. Furthermore, Figure 4(b) shows that microseismic events in the east wing goaf occurred mainly in the high rock layer, while those in the west wing goaf occurred primarily in the low rock layer, indicating that the overburden fissures in the east and west wing goaf of the asymmetric isolated coal pillar area were at different heights.Figure 4
Microseismic localization map of the asymmetric isolated coal pillar area (2018.01.01∼2018.06.30): (a) microseismic positioning plan; (b) microseismic positioning orientation profile.
(a)(b)With a precise picture of the roadway response in the coal pillar area and microseismic monitoring, the asymmetric isolated coal pillar in the mining area was affected by the coal seam mining on both sides. Furthermore, the stress concentration of the roadway surrounding the rock system was high, the overlying rock was highly active, and the roof's stability was unsatisfactory. Therefore, to reduce the safety of the roadway group in the isolated coal pillar area during the design working face excavation and later working face retrieval, it was imperative to study the mechanism of coal burst in the asymmetric isolated coal pillar area.


## 3. Stress Distribution Characteristics of Coal Seams in Asymmetric Isolated Coal Pillar Area

### 3.1. Analysis of the Spatial Structure of the Overburden Rock in Isolated Coal Pillar Area

#### 3.1.1. Analysis of the Evolution Pattern of Overburden Fracture

The critical stratum theory suggested that the overlying essential strata controlled all rock activity in the overlying rock layer of the quarry either locally or up to the surface [17]. Meanwhile, the maximum surface subsidence value was related to the mining depth H and L. Therefore, when the L/H value was more significant, it indicated a greater mining degree. Therefore, when the size L of the mining area reached or exceeded (1.2∼1.4) H, the surface achieved the full mining degree, and the surface subsidence coefficient did not increase any further [19].Based on the current mining status of the working face in the central mining district, the roof collapse of the east wing of the mining area was sufficient for complete extraction. However, the overlying rock fissure development height of the west wing of the mining area was limited, and the roof collapse was not enough, leading to a mining insufficiency (see Table1).Table 1
Judgment table for district’s mining adoption level.
PositionGoaf length (m)Goaf width (m)Mining depth (m)L/HLevel of adoptionEast wing goaf161011007501.47Full extractionWest wing goaf (I-I section)6503100.41Insufficiency miningWest wing goaf (III-III section)7001800.24Meanwhile, the general rock movement theory suggested that the structure’s maximum height formed by the fractured rock layer above the quarry during the mining insufficiency phase was one-half of the mined-out area [20]. Then, in the overburden strike profile, the isolated coal pillar I-I section area was an asymmetric “T”-type isolated structure with one entire side, complete extraction, and the other mining insufficiency (the overburden breakage height was about 150 m). Finally, the III-III section area was an asymmetric “T”-shaped isolated structure with one complete side extraction and one side insufficiency mining (the overburden breakage height was about 80 m).Surface subsidence observation points were established before mining on the west wing of the central mining district, and the measured surface subsidence curves are in Figure5. According to the surface subsidence monitoring data for October 31, 2018, the surface subsidence from measurement point L67 to D82 on the surface was 0 to 800 mm using the integral probability method [21].`(3)Wmax=qmcosβ,where q is the surface submersion coefficient and takes 0.4∼0.95; m is the coal seam thickness, m = 5.47 m; and β is the dip angle of the coal seam, β = 8°.Figure 5
Surface subsidence curve.Equation (3) yields the empirical range for the maximum ground subsidence on the west wing of the central mining district to be 2,160 mm < Wmax < 5,140 mm. Since this ground subsidence on the west wing of the mining area was smaller than the practical minimum subsidence for this area, we inferred that the overlying high essential stratum in the west wing goaf was intact.Therefore, in the overburden strike profile, the asymmetric isolated coal pillar area in the central mining district was an asymmetric “T”-type isolated coal pillar with one side thoroughly mined and another not fully excavated. In addition, due to the fault-protected coal pillar in the west wing goaf, the overburden breakage height differed on both sides of the fault-protected coal pillar (see Figure6).Figure 6
T-type overlying structure of isolated pillar in central mining district: (a) I-I section overburden structure; (b) III-III section overburden structure.
(a)(b)


#### 3.1.2. Analysis of Overburden Movement Process

For the asymmetric “T”-type isolated structure in the deep asymmetric isolated coal pillar area, the roof stratum breaking motion process was included [22]: low-stand low key strata tension breaking ⟶ low-stand critical block rotary extrusion ⟶ rotational instability of low-stand low key strata of “T”-type long wall and short wall low-stand typical key strata shear slip instability ⟶ high-stand low key strata tension breaking ⟶ high-stand key block rotary extrusion ⟶ low-stand strata passive instability ⟶ breaking of primary key strata of “T”-type long wall. The process of local overburden breaking movement in the isolated coal pillar area is shown schematically in Figure 7.Figure 7
The process of mining earthquake induced by an overburden fracture.Therefore, the breaking of critical layers at all levels above the goaf on both sides of the asymmetric isolated coal pillar area, the synergistic movement of overburden rock under significant goaf conditions, and the change of spatial structure will not only cause the increase of stress in the coal and rock layer in the asymmetric isolated coal pillar area. However, also the vibration generated by the breaking of overburden rock will impose a dynamic stress load on the coal and rock layer in the coal pillar area in the form of stress waves, which prompt the destabilization of the surrounding rock system of the roadway in the asymmetric coal pillar area that induced the dynamic coal burst disaster.


### 3.2. Stress Distribution Characterization in the Isolated Coal Pillar Area

The stress distribution pattern on the isolated coal pillar area in the central mining district varied with width dimension,D of the protected coal pillar, and the influence range S of the oversupporting pressure at the working face. However, there were four different forms [23], as shown in Figure 8.Figure 8
Distribution of superimposed stress field in isolated coal pillar area: (a) double peak; (b) saddle-shaped; (c) platform-shaped; (d) single peak.
(a)(b)(c)(d)When the width of the isolated coal pillar was large (D > 2S), the coal pillar support capacity was large, and the peak stress was located on both sides of the coal pillar and not superimposed. The central stress of the coal pillar was the original rock stress. At this instance, the supporting pressure distribution on the coal pillar area was a “double-peak” distribution (see Figure 8(a)). When the width of the isolated coal pillar complied with 2S  >  D > S, the superimposed support pressure of the working face, the plastic zone of the coal layer on both sides of the hollow mining area gradually extended to the center of the coal pillar. As a result, the stress in the central region of the coal pillar continued to rise and was more significant than the original rock stress but still less than the peak stress. Because the peak support pressure was not yet superimposed, its peak size had not changed. At this instant, the support pressure curve was a “saddle-shaped” distribution (see Figure 8(b)). When the width of the isolated coal pillar was small (D < S), the peak support pressure at the edge of the coal column was superimposed. As a result, the uniformly distributed stress value increased in the central region of the coal pillar. The elastic core region was in the external high-stress state for a long time to produce plastic damage. At this time, the support pressure curve was a “platform-shaped” distribution (see Figure 8(c)); for a minimum width (D << S) of the isolated coal pillar, the plastic zone on both sides of the column was through. The pressure value in the central region of the column is greater than the ultimate strength of the coal seam. At this time, the pillar had no elastic core area, and destabilization was very easy to occur when disturbed by external forces. At this time, the support pressure curve was a “single peak” distribution (see Figure 8(d)).Based on the historical mining data of the working face in the central mining district, the influence range of the oversupport pressure caused during the active face recovery was generally 80 m. Hence for asymmetrically isolated coal pillars in the central mining district, the width of the pillar was more extensive than 160 m, and the stress of the pillar was characterized by a “double peak” distribution. When the width of the pillar was greater than 80 m and less than 160 m, the stress of the pillar was a “saddle-shaped” distribution; when the width of the pillar was less than 80 m, the stress of the pillar was “platform-shaped” or a “single-peak” distribution. Then, the stress distribution of the coal seam in the asymmetric isolated coal pillar area in the central mining district was initially determined as shown in Figure9, and the stress pattern in the coal pillar area was mostly a “saddle-shaped” distribution, and the local area was a “double-peak” distribution.Figure 9
Schematic diagram of stress field distribution characteristics in the isolated coal pillar area.


### 3.3. Distribution Law of Lateral Abutment Pressure in Isolated Coal Pillar Area

The analysis above of stress distribution characteristics in the isolated coal pillar area depended on the relationship between the coal pillar sizeD and the influence range S of oversupport pressure at the working face. In contrast, the spatial overburden structure in the isolated coal pillar determined the stress transfer mechanism from its overburden to the coal layer. Moreover, the different spatial overburden structures directly led to the further stress distribution in the coal pillar.As shown in Figure10, the static stress load σj in the coal seam around the coal pillar area was mainly composed of the self-weight stress σg in the surrounding rock and the stress increments Δσ1, and Δσ2 transferred to the isolated coal pillar area from the overlying rock layer in the mining area on both sides [24]. Therefore, the support pressure in the isolated coal pillar area is expressed by the following equation :(4)σj=σg+Δσ1+Δσ2,where the expression of the self-gravity stress σg is as follows:(5)σg=γH,where γ is the average unit weight of the overlying rock layer, N·m−3, and H is the average depth of the isolated coal pillar, m.Figure 10
Mechanical model of overlying rock structure.For practical engineering simplicity in calculating the stress transferred from each overburdened critical layer to the coal seam at the working face, its stress is assumed to be an isosceles triangle distribution. Then, the stress increment generated by theith critical layer was expressed through Equation (6) [24]:(6)Δσ1,2=ΔσmaxxtanαHi,0,Hitanα,Δσmax2−xtanαHi,Hitanα,2Hitanα,0,2Hitanα,+∞,where α is the angle of strata movement (°); Hi is the height of destressed zone, m; and Δσmax is the maximum value of the stress increment generated by the overburden of the ith layer of the unilateral mining area on the solitary coal pillar seam, Pa. The equation for calculating Δσmax was derived from the load Qi transferred to the isolated working face from the fundamental rock block of the ith layer in the adjacent mining area [17].(7)Qi=qi2=γLiMi+mi2,where Qi is the self-weight load of the ith key strata and the overlying rock strata it controls, MPa; Li is the length of the ith key strata, m; Mi and mi are the ith key strata thicknesses and the overlying rock layer it controls, m. So then, the stress increment transferred from the overburden layer to the isolated coal pillar mainly depended on the overhanging length of the critical strata, the evolution height of the overburden, and the layer thickness.Based on the geological data of the asymmetric isolated coal pillar area in the central mining district, the mechanical parameters required to calculate stress distribution in the coal pillar area were as follows in order (see Table2).Table 2
Overburden parameters for breaking information.
iHi (m)Li (m)Mi (m)mi (m)σi (MPa)α (°)γ (kN·m−3)130.1535.226.3433.367.248325.5388.450.823.2858.624.892167.9465.718.6878.523.954271.0988.630.57130.635.475452.67134.271.3432512.21Section II-II’s support pressure distribution curve in the isolated coal pillar area was obtained (see Figure11(a)).Figure 11
Schematic diagram of supporting pressure distribution in isolated coal pillar area: (a) II-II section; (b) III-III section.
(a)(b)From the support pressure distribution on section II-II, it was evident that the overall asymmetric “saddle-shaped” distribution of coal pillar support pressure was affected by the asymmetric “T”-shaped spatial overburden structure in the isolated coal pillar area. Furthermore, the peak support pressure on the east wing goaf was the highest, with a theoretically calculated value of about 33.5 MPa, about 34 m from the edge of the mining area.Similarly, the overall asymmetric “saddle-shaped” distribution of coal pillar support pressure in the I-I section and IV-IV section, respectively, was theoretically calculated. The overall asymmetric “double-peak” distribution of coal pillar support pressure in the III-III section, with the peak support pressure located on the side of the east wing goaf, is shown in Figure11(b).


## 3.1. Analysis of the Spatial Structure of the Overburden Rock in Isolated Coal Pillar Area

### 3.1.1. Analysis of the Evolution Pattern of Overburden Fracture

The critical stratum theory suggested that the overlying essential strata controlled all rock activity in the overlying rock layer of the quarry either locally or up to the surface [17]. Meanwhile, the maximum surface subsidence value was related to the mining depth H and L. Therefore, when the L/H value was more significant, it indicated a greater mining degree. Therefore, when the size L of the mining area reached or exceeded (1.2∼1.4) H, the surface achieved the full mining degree, and the surface subsidence coefficient did not increase any further [19].Based on the current mining status of the working face in the central mining district, the roof collapse of the east wing of the mining area was sufficient for complete extraction. However, the overlying rock fissure development height of the west wing of the mining area was limited, and the roof collapse was not enough, leading to a mining insufficiency (see Table1).Table 1
Judgment table for district’s mining adoption level.
PositionGoaf length (m)Goaf width (m)Mining depth (m)L/HLevel of adoptionEast wing goaf161011007501.47Full extractionWest wing goaf (I-I section)6503100.41Insufficiency miningWest wing goaf (III-III section)7001800.24Meanwhile, the general rock movement theory suggested that the structure’s maximum height formed by the fractured rock layer above the quarry during the mining insufficiency phase was one-half of the mined-out area [20]. Then, in the overburden strike profile, the isolated coal pillar I-I section area was an asymmetric “T”-type isolated structure with one entire side, complete extraction, and the other mining insufficiency (the overburden breakage height was about 150 m). Finally, the III-III section area was an asymmetric “T”-shaped isolated structure with one complete side extraction and one side insufficiency mining (the overburden breakage height was about 80 m).Surface subsidence observation points were established before mining on the west wing of the central mining district, and the measured surface subsidence curves are in Figure5. According to the surface subsidence monitoring data for October 31, 2018, the surface subsidence from measurement point L67 to D82 on the surface was 0 to 800 mm using the integral probability method [21].`(3)Wmax=qmcosβ,where q is the surface submersion coefficient and takes 0.4∼0.95; m is the coal seam thickness, m = 5.47 m; and β is the dip angle of the coal seam, β = 8°.Figure 5
Surface subsidence curve.Equation (3) yields the empirical range for the maximum ground subsidence on the west wing of the central mining district to be 2,160 mm < Wmax < 5,140 mm. Since this ground subsidence on the west wing of the mining area was smaller than the practical minimum subsidence for this area, we inferred that the overlying high essential stratum in the west wing goaf was intact.Therefore, in the overburden strike profile, the asymmetric isolated coal pillar area in the central mining district was an asymmetric “T”-type isolated coal pillar with one side thoroughly mined and another not fully excavated. In addition, due to the fault-protected coal pillar in the west wing goaf, the overburden breakage height differed on both sides of the fault-protected coal pillar (see Figure6).Figure 6
T-type overlying structure of isolated pillar in central mining district: (a) I-I section overburden structure; (b) III-III section overburden structure.
(a)(b)


### 3.1.2. Analysis of Overburden Movement Process

For the asymmetric “T”-type isolated structure in the deep asymmetric isolated coal pillar area, the roof stratum breaking motion process was included [22]: low-stand low key strata tension breaking ⟶ low-stand critical block rotary extrusion ⟶ rotational instability of low-stand low key strata of “T”-type long wall and short wall low-stand typical key strata shear slip instability ⟶ high-stand low key strata tension breaking ⟶ high-stand key block rotary extrusion ⟶ low-stand strata passive instability ⟶ breaking of primary key strata of “T”-type long wall. The process of local overburden breaking movement in the isolated coal pillar area is shown schematically in Figure 7.Figure 7
The process of mining earthquake induced by an overburden fracture.Therefore, the breaking of critical layers at all levels above the goaf on both sides of the asymmetric isolated coal pillar area, the synergistic movement of overburden rock under significant goaf conditions, and the change of spatial structure will not only cause the increase of stress in the coal and rock layer in the asymmetric isolated coal pillar area. However, also the vibration generated by the breaking of overburden rock will impose a dynamic stress load on the coal and rock layer in the coal pillar area in the form of stress waves, which prompt the destabilization of the surrounding rock system of the roadway in the asymmetric coal pillar area that induced the dynamic coal burst disaster.


## 3.1.1. Analysis of the Evolution Pattern of Overburden Fracture

The critical stratum theory suggested that the overlying essential strata controlled all rock activity in the overlying rock layer of the quarry either locally or up to the surface [17]. Meanwhile, the maximum surface subsidence value was related to the mining depth H and L. Therefore, when the L/H value was more significant, it indicated a greater mining degree. Therefore, when the size L of the mining area reached or exceeded (1.2∼1.4) H, the surface achieved the full mining degree, and the surface subsidence coefficient did not increase any further [19].Based on the current mining status of the working face in the central mining district, the roof collapse of the east wing of the mining area was sufficient for complete extraction. However, the overlying rock fissure development height of the west wing of the mining area was limited, and the roof collapse was not enough, leading to a mining insufficiency (see Table1).Table 1
Judgment table for district’s mining adoption level.
PositionGoaf length (m)Goaf width (m)Mining depth (m)L/HLevel of adoptionEast wing goaf161011007501.47Full extractionWest wing goaf (I-I section)6503100.41Insufficiency miningWest wing goaf (III-III section)7001800.24Meanwhile, the general rock movement theory suggested that the structure’s maximum height formed by the fractured rock layer above the quarry during the mining insufficiency phase was one-half of the mined-out area [20]. Then, in the overburden strike profile, the isolated coal pillar I-I section area was an asymmetric “T”-type isolated structure with one entire side, complete extraction, and the other mining insufficiency (the overburden breakage height was about 150 m). Finally, the III-III section area was an asymmetric “T”-shaped isolated structure with one complete side extraction and one side insufficiency mining (the overburden breakage height was about 80 m).Surface subsidence observation points were established before mining on the west wing of the central mining district, and the measured surface subsidence curves are in Figure5. According to the surface subsidence monitoring data for October 31, 2018, the surface subsidence from measurement point L67 to D82 on the surface was 0 to 800 mm using the integral probability method [21].`(3)Wmax=qmcosβ,where q is the surface submersion coefficient and takes 0.4∼0.95; m is the coal seam thickness, m = 5.47 m; and β is the dip angle of the coal seam, β = 8°.Figure 5
Surface subsidence curve.Equation (3) yields the empirical range for the maximum ground subsidence on the west wing of the central mining district to be 2,160 mm < Wmax < 5,140 mm. Since this ground subsidence on the west wing of the mining area was smaller than the practical minimum subsidence for this area, we inferred that the overlying high essential stratum in the west wing goaf was intact.Therefore, in the overburden strike profile, the asymmetric isolated coal pillar area in the central mining district was an asymmetric “T”-type isolated coal pillar with one side thoroughly mined and another not fully excavated. In addition, due to the fault-protected coal pillar in the west wing goaf, the overburden breakage height differed on both sides of the fault-protected coal pillar (see Figure6).Figure 6
T-type overlying structure of isolated pillar in central mining district: (a) I-I section overburden structure; (b) III-III section overburden structure.
(a)(b)


## 3.1.2. Analysis of Overburden Movement Process

For the asymmetric “T”-type isolated structure in the deep asymmetric isolated coal pillar area, the roof stratum breaking motion process was included [22]: low-stand low key strata tension breaking ⟶ low-stand critical block rotary extrusion ⟶ rotational instability of low-stand low key strata of “T”-type long wall and short wall low-stand typical key strata shear slip instability ⟶ high-stand low key strata tension breaking ⟶ high-stand key block rotary extrusion ⟶ low-stand strata passive instability ⟶ breaking of primary key strata of “T”-type long wall. The process of local overburden breaking movement in the isolated coal pillar area is shown schematically in Figure 7.Figure 7
The process of mining earthquake induced by an overburden fracture.Therefore, the breaking of critical layers at all levels above the goaf on both sides of the asymmetric isolated coal pillar area, the synergistic movement of overburden rock under significant goaf conditions, and the change of spatial structure will not only cause the increase of stress in the coal and rock layer in the asymmetric isolated coal pillar area. However, also the vibration generated by the breaking of overburden rock will impose a dynamic stress load on the coal and rock layer in the coal pillar area in the form of stress waves, which prompt the destabilization of the surrounding rock system of the roadway in the asymmetric coal pillar area that induced the dynamic coal burst disaster.


## 3.2. Stress Distribution Characterization in the Isolated Coal Pillar Area

The stress distribution pattern on the isolated coal pillar area in the central mining district varied with width dimension,D of the protected coal pillar, and the influence range S of the oversupporting pressure at the working face. However, there were four different forms [23], as shown in Figure 8.Figure 8
Distribution of superimposed stress field in isolated coal pillar area: (a) double peak; (b) saddle-shaped; (c) platform-shaped; (d) single peak.
(a)(b)(c)(d)When the width of the isolated coal pillar was large (D > 2S), the coal pillar support capacity was large, and the peak stress was located on both sides of the coal pillar and not superimposed. The central stress of the coal pillar was the original rock stress. At this instance, the supporting pressure distribution on the coal pillar area was a “double-peak” distribution (see Figure 8(a)). When the width of the isolated coal pillar complied with 2S  >  D > S, the superimposed support pressure of the working face, the plastic zone of the coal layer on both sides of the hollow mining area gradually extended to the center of the coal pillar. As a result, the stress in the central region of the coal pillar continued to rise and was more significant than the original rock stress but still less than the peak stress. Because the peak support pressure was not yet superimposed, its peak size had not changed. At this instant, the support pressure curve was a “saddle-shaped” distribution (see Figure 8(b)). When the width of the isolated coal pillar was small (D < S), the peak support pressure at the edge of the coal column was superimposed. As a result, the uniformly distributed stress value increased in the central region of the coal pillar. The elastic core region was in the external high-stress state for a long time to produce plastic damage. At this time, the support pressure curve was a “platform-shaped” distribution (see Figure 8(c)); for a minimum width (D << S) of the isolated coal pillar, the plastic zone on both sides of the column was through. The pressure value in the central region of the column is greater than the ultimate strength of the coal seam. At this time, the pillar had no elastic core area, and destabilization was very easy to occur when disturbed by external forces. At this time, the support pressure curve was a “single peak” distribution (see Figure 8(d)).Based on the historical mining data of the working face in the central mining district, the influence range of the oversupport pressure caused during the active face recovery was generally 80 m. Hence for asymmetrically isolated coal pillars in the central mining district, the width of the pillar was more extensive than 160 m, and the stress of the pillar was characterized by a “double peak” distribution. When the width of the pillar was greater than 80 m and less than 160 m, the stress of the pillar was a “saddle-shaped” distribution; when the width of the pillar was less than 80 m, the stress of the pillar was “platform-shaped” or a “single-peak” distribution. Then, the stress distribution of the coal seam in the asymmetric isolated coal pillar area in the central mining district was initially determined as shown in Figure9, and the stress pattern in the coal pillar area was mostly a “saddle-shaped” distribution, and the local area was a “double-peak” distribution.Figure 9
Schematic diagram of stress field distribution characteristics in the isolated coal pillar area.


## 3.3. Distribution Law of Lateral Abutment Pressure in Isolated Coal Pillar Area

The analysis above of stress distribution characteristics in the isolated coal pillar area depended on the relationship between the coal pillar sizeD and the influence range S of oversupport pressure at the working face. In contrast, the spatial overburden structure in the isolated coal pillar determined the stress transfer mechanism from its overburden to the coal layer. Moreover, the different spatial overburden structures directly led to the further stress distribution in the coal pillar.As shown in Figure10, the static stress load σj in the coal seam around the coal pillar area was mainly composed of the self-weight stress σg in the surrounding rock and the stress increments Δσ1, and Δσ2 transferred to the isolated coal pillar area from the overlying rock layer in the mining area on both sides [24]. Therefore, the support pressure in the isolated coal pillar area is expressed by the following equation :(4)σj=σg+Δσ1+Δσ2,where the expression of the self-gravity stress σg is as follows:(5)σg=γH,where γ is the average unit weight of the overlying rock layer, N·m−3, and H is the average depth of the isolated coal pillar, m.Figure 10
Mechanical model of overlying rock structure.For practical engineering simplicity in calculating the stress transferred from each overburdened critical layer to the coal seam at the working face, its stress is assumed to be an isosceles triangle distribution. Then, the stress increment generated by theith critical layer was expressed through Equation (6) [24]:(6)Δσ1,2=ΔσmaxxtanαHi,0,Hitanα,Δσmax2−xtanαHi,Hitanα,2Hitanα,0,2Hitanα,+∞,where α is the angle of strata movement (°); Hi is the height of destressed zone, m; and Δσmax is the maximum value of the stress increment generated by the overburden of the ith layer of the unilateral mining area on the solitary coal pillar seam, Pa. The equation for calculating Δσmax was derived from the load Qi transferred to the isolated working face from the fundamental rock block of the ith layer in the adjacent mining area [17].(7)Qi=qi2=γLiMi+mi2,where Qi is the self-weight load of the ith key strata and the overlying rock strata it controls, MPa; Li is the length of the ith key strata, m; Mi and mi are the ith key strata thicknesses and the overlying rock layer it controls, m. So then, the stress increment transferred from the overburden layer to the isolated coal pillar mainly depended on the overhanging length of the critical strata, the evolution height of the overburden, and the layer thickness.Based on the geological data of the asymmetric isolated coal pillar area in the central mining district, the mechanical parameters required to calculate stress distribution in the coal pillar area were as follows in order (see Table2).Table 2
Overburden parameters for breaking information.
iHi (m)Li (m)Mi (m)mi (m)σi (MPa)α (°)γ (kN·m−3)130.1535.226.3433.367.248325.5388.450.823.2858.624.892167.9465.718.6878.523.954271.0988.630.57130.635.475452.67134.271.3432512.21Section II-II’s support pressure distribution curve in the isolated coal pillar area was obtained (see Figure11(a)).Figure 11
Schematic diagram of supporting pressure distribution in isolated coal pillar area: (a) II-II section; (b) III-III section.
(a)(b)From the support pressure distribution on section II-II, it was evident that the overall asymmetric “saddle-shaped” distribution of coal pillar support pressure was affected by the asymmetric “T”-shaped spatial overburden structure in the isolated coal pillar area. Furthermore, the peak support pressure on the east wing goaf was the highest, with a theoretically calculated value of about 33.5 MPa, about 34 m from the edge of the mining area.Similarly, the overall asymmetric “saddle-shaped” distribution of coal pillar support pressure in the I-I section and IV-IV section, respectively, was theoretically calculated. The overall asymmetric “double-peak” distribution of coal pillar support pressure in the III-III section, with the peak support pressure located on the side of the east wing goaf, is shown in Figure11(b).


## 4. Numerical Simulation Analysis of Stress Field in Asymmetric Isolated Coal Pillar Area

### 4.1. Mining District Modeling

The FLAC3D numerical model was established to study the stress distribution characteristics of the coal and rock layer in the asymmetric isolated coal column area after the recovery of the working face in the central mining district. After simplifying the actual geological conditions in the mining area, the model size was 3,900 × 1,800 × 400 m (X, Y, Z) (see Figure 12). The horizontal and vertical displacement constraints were applied to the bottom at both model ends. As a result, the model's top was a free surface, and a 10 MPa uniform load replaced the overlying rock layer to simulate a 400 m depth. According to the mining practice, the relevant mechanical parameters determined through field geological investigation and relevant parameters obtained from rock mechanics testing were considered in the simulation calculation. The mechanical parameters used in this numerical modeling are in Table 3.Figure 12
Central mining district model diagram.Table 3
Main coal and rock mechanics parameters.
LithologyDensity (kg·m−3)Bulk modulus (GPa)Shear modulus (GPa)Internal friction angle (°)Cohesion (MPa)Tensile strength (MPa)7#Coal1 4101.90.9281.22.30Fine sandstone2 61111.26.034710.40Siltstone2 5897.13.53685.04Medium-grained sandstone2 66110.25.03588.15Kern stone2 6368.04.03296.84Mudstone2 5805.82.83342.83Sandy mudstone2 5906.32.93253.35


### 4.2. Characteristics of Static Load Stress Field Distribution in the Isolated Coal Pillar Area

The working faces were excavated following the actual mining sequence in the central mining district. The vertical stress distribution in the isolated coal pillar area is in Figure13. Affected by the wide range of hollowing in the central mining district and the superposition of oversupport pressure on both sides in the isolated coal pillar area, the overall stress level in the coal and rock layer in the coal pillar area rises, and the stress concentration is high.Figure 13
Stress distribution map.Figure14(a) shows the vertical stress distribution curve of the coal seam under the directional profile of the isolated coal pillar area. It was evident from the figure that the stress distribution in the isolated coal pillar area is an asymmetric “saddle-shaped” distribution. The peak stress was on the side of the east wing goaf due to the different spatial mining scale and overlying rock structure on both sides of the coal pillar area, causing the stress in the eastern part of the coal pillar area to be higher than that in the western region, coupled with the presence of faults in the west wing goaf to protect the coal pillar and play a specific bearing role on the roof. On the roadway layout area, the overall stress value of the coal seam showed the trend of track roadway > return roadway > haulage roadway because the return roadway is arranged in the coal seam. The simulation results showed that the stress in the roadway area decreases, but the lateral support pressure affects the site in different ranges in the roadway sides.Figure 14
Stress change curve of coal and rock mass in isolated coal pillar area: (a) coal seam stress distribution in the coal pillar area (strike profile); (b) stress distribution in track roadway roof (tendency profile); (c) stress distribution in return roadway roof (tendency profile).
(a)(b)(c)Figures14(b) and 14(c) show the vertical stress and stress increment distribution of the rock layer on the roof of the track and return roadway in the isolated coal pillar area, respectively. The simulation results showed that with the continuous mining of the working face in the mining district, the high-stress area is mainly concentrated in the area of 200∼1100 m in the coal pillar area, and the peak stress was at 400∼600 m in the coal pillar area, which coincided with the coal burst revealed site in 2016. As a result, the peak vertical stress in the roof of the track roadway increased from 19.25 MPa to 47.6 MPa before mining, with a stress increment of 147%; the peak vertical stress in the return roadway increased from 19.25 MPa to 33.4 MPa before mining. With a stress increment of 73.5%, the peak vertical stress in the isolated coal pillar area continued to increase with the continuous rise of the mining range, the peak stress continued to grow, and the stress concentration coefficient increased. As a result, the stress concentration coefficient increased significantly. At the same time, a second stress peak area appears in the stress distribution in the roof of the track roadway, indicating that the stress shifts to the deeper part of the mining district with the increase of the mining area.The numerical simulation results showed that the stress concentration in the coal and rock layer was high in asymmetric isolated coal pillar areas. As a result, the stress peak in the pillar’s middle increased. As a result, the stress shifted to the deep part of the mining district. Furthermore, the stress in the coal seam resulted in an asymmetric “saddle shape.” Due to the unreasonable size of the asymmetric isolated coal pillar area, forming a “high stress serrated isolated coal pillar” made the isolated coal pillar area reach the critical state of overall instability impact or unstable failure state.


### 4.3. Characterization of Dynamic Load Response of Roadway in Isolated Coal Pillar Area

The dynamic load action was applied to the model based on the simulation results for the central mining district under the static load of the asymmetric isolated coal pillar area. Considering that the arranged return roadway was in the coal seam, the monitoring points were set on the roof and floor layers, and the left and right sides of the roadway are seen in the red star area of Figure1 to monitor the dynamic response characteristics during the dynamic load action. The vibration frequency of the stress wave was 20 Hz, the simulated dynamic load time was two cycles (0.1 s), and the action time was 0.3 s, containing the dynamic load from generation to disappearance. The model mechanical damping was selected as local damping with a coefficient of 0.1571 and set the static boundary.According to the structural distribution characteristics of study results of the overburden rock in isolated coal pillar, the location of the earthquake source was set in the overburden rock layer of the west wing of the coal pillar III-III section, and the vertical distance of the earthquake source was about 90 m from the coal seam. Therefore, the horizontal distance from the return roadway was almost 100 m. A dynamic load simulation scheme with a specific location for the earthquake source, a 50 MPa dynamic load, was applied to the balanced static load model to monitor the surrounding rock’s maximum principal stress evolution characteristics in the return roadway [25].Figure15 shows the evolution of the maximum principal stress in the roadway under the condition of 50 MPa dynamic load. Overall, the maximum principal stress at each measurement point characterized that the roadway right side > roadway left side > roadway floor > roadway roof, endorsing that the principal stress is asymmetrically distributed in the rock surrounding the roadway. According to the change rule for the maximum principal stress in the roadway roof, floor, and sides in dynamic load, the curve was divisible into three stages. In the first stage (0∼0.02 s), the dynamic stress wave had not spread to the monitoring point, and the stress at each monitoring point was the sole result of the static load. Stage II (0.02∼0.175 s) process was the active dynamic stage, in that the dynamic stress wave arrived the moment the maximum principal stress in the rock surrounding the roadway began to rise, following the significant decline in the stress wave with cyclic fluctuations and gradual decaying of the vibration amplitude. As a result, under the dynamic load disturbance, the fluctuation amplitude (stress deviation 2.60 MPa) observed on the right side of roadway > roadway left side (stress deviation 2.34 MPa) > roadway floor (stress deviation 1.79 MPa) > roadway roof (stress deviation 1.47 MPa). The characteristics of the maximum principal stress distribution in the surrounding rock conformed to the above observations. In stage III (0.175∼0.30 s), the maximum principal stress at each monitoring point was stable, and the stress variation did not exceed 0.1 MPa, indicating that the roadway was not disturbed by the dynamic loading.Figure 15
Characteristics of the dynamic load response of the roadway in the isolated coal pillar area.The dynamic load simulation results showed that when the coal seam roadway in the isolated coal column area was disturbed by dynamic load under high static load, the stress system in the rock surrounding the roadway produced more significant stress changes, and the roadway was more susceptible to destabilizing damage.


## 4.1. Mining District Modeling

The FLAC3D numerical model was established to study the stress distribution characteristics of the coal and rock layer in the asymmetric isolated coal column area after the recovery of the working face in the central mining district. After simplifying the actual geological conditions in the mining area, the model size was 3,900 × 1,800 × 400 m (X, Y, Z) (see Figure 12). The horizontal and vertical displacement constraints were applied to the bottom at both model ends. As a result, the model's top was a free surface, and a 10 MPa uniform load replaced the overlying rock layer to simulate a 400 m depth. According to the mining practice, the relevant mechanical parameters determined through field geological investigation and relevant parameters obtained from rock mechanics testing were considered in the simulation calculation. The mechanical parameters used in this numerical modeling are in Table 3.Figure 12
Central mining district model diagram.Table 3
Main coal and rock mechanics parameters.
LithologyDensity (kg·m−3)Bulk modulus (GPa)Shear modulus (GPa)Internal friction angle (°)Cohesion (MPa)Tensile strength (MPa)7#Coal1 4101.90.9281.22.30Fine sandstone2 61111.26.034710.40Siltstone2 5897.13.53685.04Medium-grained sandstone2 66110.25.03588.15Kern stone2 6368.04.03296.84Mudstone2 5805.82.83342.83Sandy mudstone2 5906.32.93253.35


## 4.2. Characteristics of Static Load Stress Field Distribution in the Isolated Coal Pillar Area

The working faces were excavated following the actual mining sequence in the central mining district. The vertical stress distribution in the isolated coal pillar area is in Figure13. Affected by the wide range of hollowing in the central mining district and the superposition of oversupport pressure on both sides in the isolated coal pillar area, the overall stress level in the coal and rock layer in the coal pillar area rises, and the stress concentration is high.Figure 13
Stress distribution map.Figure14(a) shows the vertical stress distribution curve of the coal seam under the directional profile of the isolated coal pillar area. It was evident from the figure that the stress distribution in the isolated coal pillar area is an asymmetric “saddle-shaped” distribution. The peak stress was on the side of the east wing goaf due to the different spatial mining scale and overlying rock structure on both sides of the coal pillar area, causing the stress in the eastern part of the coal pillar area to be higher than that in the western region, coupled with the presence of faults in the west wing goaf to protect the coal pillar and play a specific bearing role on the roof. On the roadway layout area, the overall stress value of the coal seam showed the trend of track roadway > return roadway > haulage roadway because the return roadway is arranged in the coal seam. The simulation results showed that the stress in the roadway area decreases, but the lateral support pressure affects the site in different ranges in the roadway sides.Figure 14
Stress change curve of coal and rock mass in isolated coal pillar area: (a) coal seam stress distribution in the coal pillar area (strike profile); (b) stress distribution in track roadway roof (tendency profile); (c) stress distribution in return roadway roof (tendency profile).
(a)(b)(c)Figures14(b) and 14(c) show the vertical stress and stress increment distribution of the rock layer on the roof of the track and return roadway in the isolated coal pillar area, respectively. The simulation results showed that with the continuous mining of the working face in the mining district, the high-stress area is mainly concentrated in the area of 200∼1100 m in the coal pillar area, and the peak stress was at 400∼600 m in the coal pillar area, which coincided with the coal burst revealed site in 2016. As a result, the peak vertical stress in the roof of the track roadway increased from 19.25 MPa to 47.6 MPa before mining, with a stress increment of 147%; the peak vertical stress in the return roadway increased from 19.25 MPa to 33.4 MPa before mining. With a stress increment of 73.5%, the peak vertical stress in the isolated coal pillar area continued to increase with the continuous rise of the mining range, the peak stress continued to grow, and the stress concentration coefficient increased. As a result, the stress concentration coefficient increased significantly. At the same time, a second stress peak area appears in the stress distribution in the roof of the track roadway, indicating that the stress shifts to the deeper part of the mining district with the increase of the mining area.The numerical simulation results showed that the stress concentration in the coal and rock layer was high in asymmetric isolated coal pillar areas. As a result, the stress peak in the pillar’s middle increased. As a result, the stress shifted to the deep part of the mining district. Furthermore, the stress in the coal seam resulted in an asymmetric “saddle shape.” Due to the unreasonable size of the asymmetric isolated coal pillar area, forming a “high stress serrated isolated coal pillar” made the isolated coal pillar area reach the critical state of overall instability impact or unstable failure state.


## 4.3. Characterization of Dynamic Load Response of Roadway in Isolated Coal Pillar Area

The dynamic load action was applied to the model based on the simulation results for the central mining district under the static load of the asymmetric isolated coal pillar area. Considering that the arranged return roadway was in the coal seam, the monitoring points were set on the roof and floor layers, and the left and right sides of the roadway are seen in the red star area of Figure1 to monitor the dynamic response characteristics during the dynamic load action. The vibration frequency of the stress wave was 20 Hz, the simulated dynamic load time was two cycles (0.1 s), and the action time was 0.3 s, containing the dynamic load from generation to disappearance. The model mechanical damping was selected as local damping with a coefficient of 0.1571 and set the static boundary.According to the structural distribution characteristics of study results of the overburden rock in isolated coal pillar, the location of the earthquake source was set in the overburden rock layer of the west wing of the coal pillar III-III section, and the vertical distance of the earthquake source was about 90 m from the coal seam. Therefore, the horizontal distance from the return roadway was almost 100 m. A dynamic load simulation scheme with a specific location for the earthquake source, a 50 MPa dynamic load, was applied to the balanced static load model to monitor the surrounding rock’s maximum principal stress evolution characteristics in the return roadway [25].Figure15 shows the evolution of the maximum principal stress in the roadway under the condition of 50 MPa dynamic load. Overall, the maximum principal stress at each measurement point characterized that the roadway right side > roadway left side > roadway floor > roadway roof, endorsing that the principal stress is asymmetrically distributed in the rock surrounding the roadway. According to the change rule for the maximum principal stress in the roadway roof, floor, and sides in dynamic load, the curve was divisible into three stages. In the first stage (0∼0.02 s), the dynamic stress wave had not spread to the monitoring point, and the stress at each monitoring point was the sole result of the static load. Stage II (0.02∼0.175 s) process was the active dynamic stage, in that the dynamic stress wave arrived the moment the maximum principal stress in the rock surrounding the roadway began to rise, following the significant decline in the stress wave with cyclic fluctuations and gradual decaying of the vibration amplitude. As a result, under the dynamic load disturbance, the fluctuation amplitude (stress deviation 2.60 MPa) observed on the right side of roadway > roadway left side (stress deviation 2.34 MPa) > roadway floor (stress deviation 1.79 MPa) > roadway roof (stress deviation 1.47 MPa). The characteristics of the maximum principal stress distribution in the surrounding rock conformed to the above observations. In stage III (0.175∼0.30 s), the maximum principal stress at each monitoring point was stable, and the stress variation did not exceed 0.1 MPa, indicating that the roadway was not disturbed by the dynamic loading.Figure 15
Characteristics of the dynamic load response of the roadway in the isolated coal pillar area.The dynamic load simulation results showed that when the coal seam roadway in the isolated coal column area was disturbed by dynamic load under high static load, the stress system in the rock surrounding the roadway produced more significant stress changes, and the roadway was more susceptible to destabilizing damage.


## 5. Mechanism of Coal Burst Occurrence in Asymmetric Isolated Coal Pillar Area

Coal bursts result from the combination of dynamic stress and static stress. When the dynamic and static stress in the coal and rock mass exceeded the critical stress, the coal burst will be induced [26, 27].σj + σd ≥ σBmin, where σj and σd are static and dynamic stress, respectively, in coal/rock mass induced by microseismic events; σBmin is the critical burst stress.Comprehensive analysis of the coal burst event in the mining district roadway confirmed that the main reason was a high-stress concentration in the coal and rock layers in the area that exceeded the compressive strength of the coal and rock mass. As a result, the local coal burst phenomenon occurred in the roadway. For the asymmetric isolated coal pillar area in the central mining district, the coal pillar formed a “high stress asymmetric isolated coal pillar.” Under the action of self-weight stress at significant mining depth and lateral support pressure of the coal pillar, the isolated coal pillar progressively reached the critical state for the overall instability of the coal pillar. Then, under the action of high static load, or when it was disturbed by dynamic load (overburden movement, mining activities,etc.), it finally broke through the limiting equilibrium condition for the overall instability impact on the coal pillar area, and the coal burst accident occurred (see Figure 16).Figure 16
Dynamic and static load superimposed coal burst mechanism in the asymmetric isolated coal pillar area.Summarizing the above mechanism of coal burst occurrence of asymmetric isolated coal pillar, the main influencing factors that induced the overall destabilizing impact in the deep asymmetric isolated coal pillar area were: mining depth, the width of protective coal pillar, and overburden space structure. Therefore, the mining depth belonged to the influence factor of geological conditions, and the retention of defensive coal pillar width and overburden space structure belonged to the influence factor of mining technology. The root cause for this kind of coal burst accident was the unsatisfactory size of asymmetric isolated coal pillar, which led to the formation of “high stress serrated isolated coal pillar.” The action self-weight stress and lateral supporting pressure from the coal pillar were further affected by other stress superposition, or the concentrated stress of the coal pillar exceeded the bearing capacity of the coal seam that facilitated the coal burst phenomenon in the roadway.


## 6. Coal Burst Prevention and Control Practice in Asymmetric Isolated Coal Pillar Area

### 6.1. Prevention and Treatment Programs

Based on the above analysis for the mechanism of coal burst in the deep asymmetric isolated coal pillar area, the overall instability of the isolated coal pillar was mainly related to the degree of stress concentration in the coal seam and the spatial structure of the coal seam (the overburden). Therefore, the key to preventing and controlling this type of coal burst was to reduce the concentration of coal seam support pressure in the isolated coal pillar area and reduce the dynamic load disturbance brought by the overburden breakage. In addition, strengthening the roadway support in the coal pillar area effectively prevented and controlled the coal bursts.To minimize the dynamic load generated from the roof breakage in the isolated coal pillar area caused strong disturbance to the roadway and simultaneously destroyed the integrity of the rock layer, weakened the roof structure, and reduced the length of the overhanging roof, the roof deep-hole precracking blasting cut off actively the overhanging roof and induced the early release of the elastic energy accumulated in the roof. They took the I-I further arranged III-III sections in the isolated coal pillar area as examples. As a result, several roof-deep blasting holes were arranged every 10 m in the mining district roadway [28]. Furthermore, to destroy the integrity of the coal seam in the isolated coal pillar area and reduce the degree of stress concentration in the coal seam, extensive diameter pressure relief drilling of the coal seam is adopted to weaken the coal seam and reduce the stress level of the coal seam in the coal pillar area. The above analysis of the peak stress in the coal pillar area located on the side of the east wing goaf and the track and haulage roadway was in the stress concentration area [29]. Therefore, the deep-hole arrangement was adopted on the left side of the return roadway (corresponding to the eastern part of the coal pillar). Moreover, the usual hole arrangement was adopted on the right side to reduce the stress level in the deep stress peak area of the coal pillar area, with a drill hole arrangement spacing of 2 m. The pressure relief hole arrangement is given in Figure 17, and the pressure relief parameters are in Tables 4 and 5.Figure 17
Schematic diagram of pressure relief measures: (a) I-I sections; (b) III-III sections.
(a)(b)Table 4
Parameters of pressure relief measures for I-I section.
Drilling numberDrilling depth (m)Explosive quantity (kg)Elevation angle (°)Loaded length (m)Hole sealing length (m)Drilling diameter (mm)1#4048801216632#5060701520633#6080552022634#4050801216635#5060701520636#6080552022637#55—2——1538#20—2——153Table 5
Parameters of pressure relief measures for III-III section.
Drilling numberDrilling depth (m)Explosive quantity (kg)Elevation angle (°)Loaded length (m)Hole sealing length (m)Drilling diameter (mm)1#4560751516632#5580602020633#65100452522634#4460751516635#5580602020636#6588502222637#60—2——1538#20—2——153


### 6.2. Prevention and Treatment Effect Test

#### 6.2.1. Comparison of Microseismic Monitoring Data

The asymmetric isolated coal pillar area was subjected to multiple rounds of deep-hole blasting off the roof and pressure relief measures of coal seam drilling in the III-III section area from August 1 to August 20, 2018. The microseismic data from July 10 to July 29, 2018 (before unloading pressure) and August 1 to August 20, 2018 (after unloading) were selected for analysis, as shown in Figure18. After unloading pressure of the isolated coal pillar area, the daily average microseismic release energy decreased from 2.48 × 104 J before unloading to 1.48 × 104 J after unloading, with a reduction of about 40.3%; the daily average microseismic event count decreased from 45 times before unloading to 35 times, with a decrease of approximately 22.2%.Figure 18
Microseismic monitoring data analysis chart (III-III section).The isolated coal pillar area was subjected to multiple rounds of deep-hole blasting off the roof and pressure relief measures for coal seam drilling in the I-I section area from September 24 to September 29, 2018. Therefore, the microseismic data from September 5 to September 24, 2018 (before unloading pressure) and September 30 to October 18, 2018 (after unloading) were selected for analysis, as shown in Figure19. After unloading pressure in the isolated coal pillar area, the average daily microseismic release energy decreased from 2.98 × 104 J before decompression to 1.86 × 104 J after decompression, with a reduction of about 37.6%; the average daily microseismic event count decreased from 53 times before decompression to 40 times, with a decrease of approximately 24.5%.Figure 19
Microseismic monitoring data analysis chart (I-I section).The average daily frequency of microseismic events significantly reduced before and after unloading pressure on the central mining district's asymmetric isolated coal pillar area from the average daily energy release. This reduction was more significant than 20%, indicating the stress level reduced effectively in the rock system surrounding the roadway near the isolated coal pillar area after the decompression.


#### 6.2.2. Analysis of Seismic Computerized Tomography Results

The computerized seismic tomography (seismic CT) inversion evaluation index made a dynamic evaluation of the impact hazard on the working face. The higher the P-wave velocity (VP) of the coal rock layer of the working face, the larger the positive anomaly coefficient An of the wave velocity, indicating that the higher the stress level of the surrounding rock in the area, the more substantial the risk of coal burst [4, 30, 31].In asymmetric isolated coal pillar areas, coal burst risk was dominated primarily by static stress. Accordingly, Tables6 and 7 are an assessment criterion compiled for coal burst risk. Velocity anomaly An was determined by(8)An=Vp−VpaVpa,where Vp is the P-wave velocity in a particular medium and Vpa is the average velocity in the model. It was noteworthy that the zones with positive anomaly and negative anomaly were overstressed and pressure-relieved, respectively.Table 6
Relation between positive velocity anomalies and coal burst risk level.
Coal burst risk indexCoal burst risk levelPositive velocity anomaly,An (%)0None<51Weak5∼152Middle15∼253Strong>25Table 7
Relation between negative velocity anomalies and coal burst risk level.
Weakened degreeWeakening characteristicsNegative velocity anomaly,An (%)Stress reduction degree0None0∼−7.5<0.25−1Weak−7.5∼−150.25∼0.55−2Middle−15∼−250.55∼0.8−3Strong<−25>0.8In the paper, microseismic waveforms in isolated coal pillar area from July 10 to July 29, 2018 (III-III sections before pressure relief) and from August 1 to August 20, 2018 (III-III sections after pressure relief); September 5 to September 24, 2018 (I-I sections before pressure relief); and from September 30 to October 18, 2018 (I-I sections after pressure relief) were selected as seismic computerized tomography data. Furthermore, wave velocity (VP) and the positive anomaly coefficient of wave velocity (An) were used as indicators to evaluate the coal burst risk at the coal pillar area. Figures20 and 21 show the results of the seismic CT.Figure 20
Seismic CT detection results before and after decompression in the III-III section area of the isolated coal pillar. (a) Characteristics of wave velocity distribution before decompression (2018.07.10–2018.07.29). (b)Characteristics of wave velocity distribution after deconfliction (2018.08.01–2018.08.20). (c) Distribution characteristics of wave speed anomaly index (An) before decompression. (d) Distribution characteristics of wave speed anomaly index (An) after decompression.
(a)(b)(c)(d)Figure 21
Seismic CT detection results before and after decompression in the I-I section area of the isolated coal pillar. (a) Characteristics of wave velocity distribution before decompression (2018.09.05–2018.09.24). (b) Characteristics of wave velocity distribution after deconfliction (2018.09.30–2018.10.18). (c) Distribution characteristics of wave speed anomaly index (An) before decompression. (d) Distribution characteristics of wave speed anomaly index (An) after decompression.
(a)(b)(c)(d)The results of the seismic CT in Figure20 show that the high-stress area before pressure relief was distributed mainly in the area of III-III section in the isolated coal pillar area, with a peak wave velocity of about 6.01 km/s and a peak wave velocity positive anomaly index (An) of about 0.39. The coal pillar area had a high degree of stress concentration in the rock system’s roadway, which impacted coal bursts. The results showed that the peak value of wave velocity was about 3.3 km/s in the isolated coal pillar area after roof deep-hole blasting and coal seam drilling pressure were relieved. After pressure relief, the peak value of wave velocity decreased by 45.1%, and the peak value of the negative anomaly index of wave velocity was about −0.35. Thus, the roof weakened enormously, and the stress concentration in the pressure relief area decreased. The results showed that the stress concentration degree of the rock system surrounding the roadway in Section III-III of isolated coal pillar reduced after pressure relief, and the pressure relief effect was good.The results of the seismic CT in Figure21 showed that the high-stress area before pressure relief was distributed mainly in the area of the I-I section in the isolated coal pillar area, with a peak wave velocity of about 5.5 km/s and a peak wave velocity positive anomaly index (An) of about 0.38. The coal pillar area had a high-stress concentration in the rock system’s roadway, intensely impacting coal bursts. The results showed that the peak value of wave velocity was about 4.0 km/s in the isolated coal pillar area after roof deep-hole blasting and coal seam drilling pressure were relieved. After pressure relief, the peak value of wave velocity decreased by 27.3%, and the peak value of the negative anomaly index of wave velocity was about −0.17. Thus, the roof weakened enormously, and the stress concentration in the pressure relief area decreased. The results showed that the stress concentration degree of surrounding rock system of roadway in Section III-III of isolated coal pillar area reduced after pressure relief, and the pressure relief effect was good.


## 6.1. Prevention and Treatment Programs

Based on the above analysis for the mechanism of coal burst in the deep asymmetric isolated coal pillar area, the overall instability of the isolated coal pillar was mainly related to the degree of stress concentration in the coal seam and the spatial structure of the coal seam (the overburden). Therefore, the key to preventing and controlling this type of coal burst was to reduce the concentration of coal seam support pressure in the isolated coal pillar area and reduce the dynamic load disturbance brought by the overburden breakage. In addition, strengthening the roadway support in the coal pillar area effectively prevented and controlled the coal bursts.To minimize the dynamic load generated from the roof breakage in the isolated coal pillar area caused strong disturbance to the roadway and simultaneously destroyed the integrity of the rock layer, weakened the roof structure, and reduced the length of the overhanging roof, the roof deep-hole precracking blasting cut off actively the overhanging roof and induced the early release of the elastic energy accumulated in the roof. They took the I-I further arranged III-III sections in the isolated coal pillar area as examples. As a result, several roof-deep blasting holes were arranged every 10 m in the mining district roadway [28]. Furthermore, to destroy the integrity of the coal seam in the isolated coal pillar area and reduce the degree of stress concentration in the coal seam, extensive diameter pressure relief drilling of the coal seam is adopted to weaken the coal seam and reduce the stress level of the coal seam in the coal pillar area. The above analysis of the peak stress in the coal pillar area located on the side of the east wing goaf and the track and haulage roadway was in the stress concentration area [29]. Therefore, the deep-hole arrangement was adopted on the left side of the return roadway (corresponding to the eastern part of the coal pillar). Moreover, the usual hole arrangement was adopted on the right side to reduce the stress level in the deep stress peak area of the coal pillar area, with a drill hole arrangement spacing of 2 m. The pressure relief hole arrangement is given in Figure 17, and the pressure relief parameters are in Tables 4 and 5.Figure 17
Schematic diagram of pressure relief measures: (a) I-I sections; (b) III-III sections.
(a)(b)Table 4
Parameters of pressure relief measures for I-I section.
Drilling numberDrilling depth (m)Explosive quantity (kg)Elevation angle (°)Loaded length (m)Hole sealing length (m)Drilling diameter (mm)1#4048801216632#5060701520633#6080552022634#4050801216635#5060701520636#6080552022637#55—2——1538#20—2——153Table 5
Parameters of pressure relief measures for III-III section.
Drilling numberDrilling depth (m)Explosive quantity (kg)Elevation angle (°)Loaded length (m)Hole sealing length (m)Drilling diameter (mm)1#4560751516632#5580602020633#65100452522634#4460751516635#5580602020636#6588502222637#60—2——1538#20—2——153


## 6.2. Prevention and Treatment Effect Test

### 6.2.1. Comparison of Microseismic Monitoring Data

The asymmetric isolated coal pillar area was subjected to multiple rounds of deep-hole blasting off the roof and pressure relief measures of coal seam drilling in the III-III section area from August 1 to August 20, 2018. The microseismic data from July 10 to July 29, 2018 (before unloading pressure) and August 1 to August 20, 2018 (after unloading) were selected for analysis, as shown in Figure18. After unloading pressure of the isolated coal pillar area, the daily average microseismic release energy decreased from 2.48 × 104 J before unloading to 1.48 × 104 J after unloading, with a reduction of about 40.3%; the daily average microseismic event count decreased from 45 times before unloading to 35 times, with a decrease of approximately 22.2%.Figure 18
Microseismic monitoring data analysis chart (III-III section).The isolated coal pillar area was subjected to multiple rounds of deep-hole blasting off the roof and pressure relief measures for coal seam drilling in the I-I section area from September 24 to September 29, 2018. Therefore, the microseismic data from September 5 to September 24, 2018 (before unloading pressure) and September 30 to October 18, 2018 (after unloading) were selected for analysis, as shown in Figure19. After unloading pressure in the isolated coal pillar area, the average daily microseismic release energy decreased from 2.98 × 104 J before decompression to 1.86 × 104 J after decompression, with a reduction of about 37.6%; the average daily microseismic event count decreased from 53 times before decompression to 40 times, with a decrease of approximately 24.5%.Figure 19
Microseismic monitoring data analysis chart (I-I section).The average daily frequency of microseismic events significantly reduced before and after unloading pressure on the central mining district's asymmetric isolated coal pillar area from the average daily energy release. This reduction was more significant than 20%, indicating the stress level reduced effectively in the rock system surrounding the roadway near the isolated coal pillar area after the decompression.


### 6.2.2. Analysis of Seismic Computerized Tomography Results

The computerized seismic tomography (seismic CT) inversion evaluation index made a dynamic evaluation of the impact hazard on the working face. The higher the P-wave velocity (VP) of the coal rock layer of the working face, the larger the positive anomaly coefficient An of the wave velocity, indicating that the higher the stress level of the surrounding rock in the area, the more substantial the risk of coal burst [4, 30, 31].In asymmetric isolated coal pillar areas, coal burst risk was dominated primarily by static stress. Accordingly, Tables6 and 7 are an assessment criterion compiled for coal burst risk. Velocity anomaly An was determined by(8)An=Vp−VpaVpa,where Vp is the P-wave velocity in a particular medium and Vpa is the average velocity in the model. It was noteworthy that the zones with positive anomaly and negative anomaly were overstressed and pressure-relieved, respectively.Table 6
Relation between positive velocity anomalies and coal burst risk level.
Coal burst risk indexCoal burst risk levelPositive velocity anomaly,An (%)0None<51Weak5∼152Middle15∼253Strong>25Table 7
Relation between negative velocity anomalies and coal burst risk level.
Weakened degreeWeakening characteristicsNegative velocity anomaly,An (%)Stress reduction degree0None0∼−7.5<0.25−1Weak−7.5∼−150.25∼0.55−2Middle−15∼−250.55∼0.8−3Strong<−25>0.8In the paper, microseismic waveforms in isolated coal pillar area from July 10 to July 29, 2018 (III-III sections before pressure relief) and from August 1 to August 20, 2018 (III-III sections after pressure relief); September 5 to September 24, 2018 (I-I sections before pressure relief); and from September 30 to October 18, 2018 (I-I sections after pressure relief) were selected as seismic computerized tomography data. Furthermore, wave velocity (VP) and the positive anomaly coefficient of wave velocity (An) were used as indicators to evaluate the coal burst risk at the coal pillar area. Figures20 and 21 show the results of the seismic CT.Figure 20
Seismic CT detection results before and after decompression in the III-III section area of the isolated coal pillar. (a) Characteristics of wave velocity distribution before decompression (2018.07.10–2018.07.29). (b)Characteristics of wave velocity distribution after deconfliction (2018.08.01–2018.08.20). (c) Distribution characteristics of wave speed anomaly index (An) before decompression. (d) Distribution characteristics of wave speed anomaly index (An) after decompression.
(a)(b)(c)(d)Figure 21
Seismic CT detection results before and after decompression in the I-I section area of the isolated coal pillar. (a) Characteristics of wave velocity distribution before decompression (2018.09.05–2018.09.24). (b) Characteristics of wave velocity distribution after deconfliction (2018.09.30–2018.10.18). (c) Distribution characteristics of wave speed anomaly index (An) before decompression. (d) Distribution characteristics of wave speed anomaly index (An) after decompression.
(a)(b)(c)(d)The results of the seismic CT in Figure20 show that the high-stress area before pressure relief was distributed mainly in the area of III-III section in the isolated coal pillar area, with a peak wave velocity of about 6.01 km/s and a peak wave velocity positive anomaly index (An) of about 0.39. The coal pillar area had a high degree of stress concentration in the rock system’s roadway, which impacted coal bursts. The results showed that the peak value of wave velocity was about 3.3 km/s in the isolated coal pillar area after roof deep-hole blasting and coal seam drilling pressure were relieved. After pressure relief, the peak value of wave velocity decreased by 45.1%, and the peak value of the negative anomaly index of wave velocity was about −0.35. Thus, the roof weakened enormously, and the stress concentration in the pressure relief area decreased. The results showed that the stress concentration degree of the rock system surrounding the roadway in Section III-III of isolated coal pillar reduced after pressure relief, and the pressure relief effect was good.The results of the seismic CT in Figure21 showed that the high-stress area before pressure relief was distributed mainly in the area of the I-I section in the isolated coal pillar area, with a peak wave velocity of about 5.5 km/s and a peak wave velocity positive anomaly index (An) of about 0.38. The coal pillar area had a high-stress concentration in the rock system’s roadway, intensely impacting coal bursts. The results showed that the peak value of wave velocity was about 4.0 km/s in the isolated coal pillar area after roof deep-hole blasting and coal seam drilling pressure were relieved. After pressure relief, the peak value of wave velocity decreased by 27.3%, and the peak value of the negative anomaly index of wave velocity was about −0.17. Thus, the roof weakened enormously, and the stress concentration in the pressure relief area decreased. The results showed that the stress concentration degree of surrounding rock system of roadway in Section III-III of isolated coal pillar area reduced after pressure relief, and the pressure relief effect was good.


## 6.2.1. Comparison of Microseismic Monitoring Data

The asymmetric isolated coal pillar area was subjected to multiple rounds of deep-hole blasting off the roof and pressure relief measures of coal seam drilling in the III-III section area from August 1 to August 20, 2018. The microseismic data from July 10 to July 29, 2018 (before unloading pressure) and August 1 to August 20, 2018 (after unloading) were selected for analysis, as shown in Figure18. After unloading pressure of the isolated coal pillar area, the daily average microseismic release energy decreased from 2.48 × 104 J before unloading to 1.48 × 104 J after unloading, with a reduction of about 40.3%; the daily average microseismic event count decreased from 45 times before unloading to 35 times, with a decrease of approximately 22.2%.Figure 18
Microseismic monitoring data analysis chart (III-III section).The isolated coal pillar area was subjected to multiple rounds of deep-hole blasting off the roof and pressure relief measures for coal seam drilling in the I-I section area from September 24 to September 29, 2018. Therefore, the microseismic data from September 5 to September 24, 2018 (before unloading pressure) and September 30 to October 18, 2018 (after unloading) were selected for analysis, as shown in Figure19. After unloading pressure in the isolated coal pillar area, the average daily microseismic release energy decreased from 2.98 × 104 J before decompression to 1.86 × 104 J after decompression, with a reduction of about 37.6%; the average daily microseismic event count decreased from 53 times before decompression to 40 times, with a decrease of approximately 24.5%.Figure 19
Microseismic monitoring data analysis chart (I-I section).The average daily frequency of microseismic events significantly reduced before and after unloading pressure on the central mining district's asymmetric isolated coal pillar area from the average daily energy release. This reduction was more significant than 20%, indicating the stress level reduced effectively in the rock system surrounding the roadway near the isolated coal pillar area after the decompression.


## 6.2.2. Analysis of Seismic Computerized Tomography Results

The computerized seismic tomography (seismic CT) inversion evaluation index made a dynamic evaluation of the impact hazard on the working face. The higher the P-wave velocity (VP) of the coal rock layer of the working face, the larger the positive anomaly coefficient An of the wave velocity, indicating that the higher the stress level of the surrounding rock in the area, the more substantial the risk of coal burst [4, 30, 31].In asymmetric isolated coal pillar areas, coal burst risk was dominated primarily by static stress. Accordingly, Tables6 and 7 are an assessment criterion compiled for coal burst risk. Velocity anomaly An was determined by(8)An=Vp−VpaVpa,where Vp is the P-wave velocity in a particular medium and Vpa is the average velocity in the model. It was noteworthy that the zones with positive anomaly and negative anomaly were overstressed and pressure-relieved, respectively.Table 6
Relation between positive velocity anomalies and coal burst risk level.
Coal burst risk indexCoal burst risk levelPositive velocity anomaly,An (%)0None<51Weak5∼152Middle15∼253Strong>25Table 7
Relation between negative velocity anomalies and coal burst risk level.
Weakened degreeWeakening characteristicsNegative velocity anomaly,An (%)Stress reduction degree0None0∼−7.5<0.25−1Weak−7.5∼−150.25∼0.55−2Middle−15∼−250.55∼0.8−3Strong<−25>0.8In the paper, microseismic waveforms in isolated coal pillar area from July 10 to July 29, 2018 (III-III sections before pressure relief) and from August 1 to August 20, 2018 (III-III sections after pressure relief); September 5 to September 24, 2018 (I-I sections before pressure relief); and from September 30 to October 18, 2018 (I-I sections after pressure relief) were selected as seismic computerized tomography data. Furthermore, wave velocity (VP) and the positive anomaly coefficient of wave velocity (An) were used as indicators to evaluate the coal burst risk at the coal pillar area. Figures20 and 21 show the results of the seismic CT.Figure 20
Seismic CT detection results before and after decompression in the III-III section area of the isolated coal pillar. (a) Characteristics of wave velocity distribution before decompression (2018.07.10–2018.07.29). (b)Characteristics of wave velocity distribution after deconfliction (2018.08.01–2018.08.20). (c) Distribution characteristics of wave speed anomaly index (An) before decompression. (d) Distribution characteristics of wave speed anomaly index (An) after decompression.
(a)(b)(c)(d)Figure 21
Seismic CT detection results before and after decompression in the I-I section area of the isolated coal pillar. (a) Characteristics of wave velocity distribution before decompression (2018.09.05–2018.09.24). (b) Characteristics of wave velocity distribution after deconfliction (2018.09.30–2018.10.18). (c) Distribution characteristics of wave speed anomaly index (An) before decompression. (d) Distribution characteristics of wave speed anomaly index (An) after decompression.
(a)(b)(c)(d)The results of the seismic CT in Figure20 show that the high-stress area before pressure relief was distributed mainly in the area of III-III section in the isolated coal pillar area, with a peak wave velocity of about 6.01 km/s and a peak wave velocity positive anomaly index (An) of about 0.39. The coal pillar area had a high degree of stress concentration in the rock system’s roadway, which impacted coal bursts. The results showed that the peak value of wave velocity was about 3.3 km/s in the isolated coal pillar area after roof deep-hole blasting and coal seam drilling pressure were relieved. After pressure relief, the peak value of wave velocity decreased by 45.1%, and the peak value of the negative anomaly index of wave velocity was about −0.35. Thus, the roof weakened enormously, and the stress concentration in the pressure relief area decreased. The results showed that the stress concentration degree of the rock system surrounding the roadway in Section III-III of isolated coal pillar reduced after pressure relief, and the pressure relief effect was good.The results of the seismic CT in Figure21 showed that the high-stress area before pressure relief was distributed mainly in the area of the I-I section in the isolated coal pillar area, with a peak wave velocity of about 5.5 km/s and a peak wave velocity positive anomaly index (An) of about 0.38. The coal pillar area had a high-stress concentration in the rock system’s roadway, intensely impacting coal bursts. The results showed that the peak value of wave velocity was about 4.0 km/s in the isolated coal pillar area after roof deep-hole blasting and coal seam drilling pressure were relieved. After pressure relief, the peak value of wave velocity decreased by 27.3%, and the peak value of the negative anomaly index of wave velocity was about −0.17. Thus, the roof weakened enormously, and the stress concentration in the pressure relief area decreased. The results showed that the stress concentration degree of surrounding rock system of roadway in Section III-III of isolated coal pillar area reduced after pressure relief, and the pressure relief effect was good.


## 7. Conclusion

(1)
Based on the analysis of the overburden structure in the mining district, combined with the surface subsidence data, the asymmetric isolated coal pillar was an asymmetric “T”-shaped isolated structure in the strike-profile, and the overburden breaking movement and spatial structure change on the long wall side of the “T”-shaped led to the stress change of the rock system surrounding the coal pillar area.(2)
The analyzed results of stress distribution in the coal pillar area showed that the stress in the coal seam within the asymmetric isolated coal pillar area was mainly asymmetrical “saddle-shaped” distribution, with local “double peak” distribution, and the peak of the supporting pressure was located in the east wing goaf.(3)
The numerical simulation showed that the stress value on the striking profile of the coal pillar area gradually increased with the increase of the mining area and demonstrated an asymmetric “saddle-shaped” distribution. Furthermore, the peak stress on the tendency gradually increased and gradually shifted to the mining district's deep level, and the peak stress was in the area of 400∼600 m of the isolated coal pillar.(4)
The deep asymmetric isolated coal pillar area formed a “high stress asymmetric isolated coal pillar” under the action of factors such as self-weight stress of significant mining depth and lateral support pressure of coal pillar. As a result, the coal pillar was in critical instability under the action of high static load. When disturbed by dynamic load, it was easy to cause the stress system of the roadway surrounding rock in the coal pillar area to exceed the critical stress for coal burst, thus inducing coal burst accidents.(5)
The deep-hole blasting with a drilled pressure relief scheme in the coal seam was developed for the roof of the coal pillar area to minimize the mechanism of coal bursts induced based on the dynamic and static load on asymmetric isolated coal pillars. The field practice results showed that the pressure relief effect was beneficial; the prevention and control system of asymmetric isolated coal pillar areas in mining districts had a guiding significance for preventing and controlling coal bursts in mines under such conditions.



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*Source: 3751146-2021-12-29.xml*