# Reasonable Determination of Terminal Mining Lines Using the Stress Field with Seismic Wave Excitation in Deep Coalfaces

Authors: Guang-an Zhu; Huan Liu; Bo-ru Su; Qi-peng Jiang; Hai-yang Liu
Journal: Shock and Vibration
 (2021)

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## Abstract

Many field observations have shown that rock bursts occur frequently near the terminal mining line (TML) and dip coal pillar area in deep coalfaces. Taking the “7.26” rock burst in coalface 3302 in Xingcun Coal Mine as an example, the rock burst mechanism was investigated based on theoretical analysis and field observations, and a combined evaluation method using the stress field under seismic wave excitation was established to determine the reasonable TML of coalface 3302. Firstly, the static geological data revealed during roadway excavation were used for preevaluation of rock burst risk at the working face. By theoretically analyzing the stress transfer mechanism of the two types of the roof structure, the computational model of abutment pressure was established and the calculation method giving the abutment stress was proposed. Subsequently, a dynamic evaluation method that adopts microseismic and stress online monitoring system to monitor dynamic information, such as mine tremors and stress during coalface mining, was developed to define stress anomaly areas and then dynamically determine the TML. Finally, the proposed model was used to optimize the position of the TML of LW3302 in Xingcun Coal Mine; findings obtained in this study provide theoretical guidance for safe coal mining. Combined with the results of theoretical analysis (255 m), online stress monitoring (200 m), microseismic (MS) monitoring (262 m), and passive seismic velocity tomography (220–250 m), it can be finally determined that the width of the protective coal pillar for the TML of coalface 3302 should be at least 262 m.


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## Body

## 1. Introduction

At present, more attention has been paid to mineral exploitation at greater depths; for coal resources, current mining activity has reached 1500 m in depth. As more coal mines develop to deeper workings, the frequency and intensity of rock burst disasters increase, increasing the difficulty of their prediction and prevention [1–4].Many of field observations have shown that rock bursts occur frequently near the TML and dip coal pillar area in deep coal mining. For instance, rock bursts occur frequently in the roadway and chamber groups in the dip coal pillar area of No. 21 mining district in Qianqiu Coal Mine, Yima City. In the three rock bursts in the dip coal roadways in Zhangshuanglou Coal Mine, Xuzhou City, the distances between the rock burst area and the coalfaces are 354 m, 355 m, and 460 m, respectively. The rock bursts in the panel roadway groups on 16 February and 20 February 2017 in Cuimu Coal Mine, Baoji City, caused the closure of a 200 m gateway; thus it can be seen that rock bursts in deep coal mining show certain new characteristics; the range of influence of mining activities is wider, and localized rock bursts are gradually transformed into larger rock bursts; therefore, it is necessary to determine the reasonable TML, faced in the later mining stage of district or panel in deep coal mining.Generally, there are three methods available to estimate the TML design: similarity simulation, numerical simulation, andin situ measurement. In the similarity simulation method, Xu et al. [5] conducted a similarity simulation experiment to evaluate the influence of overlying strata instability on roadway deformation and failure, and the most reasonable TML was determined. Zhou et al. [6] performed a model test of the disturbing characteristics below roadway with the axis perpendicular to the excavation orientation during the excavation of the upper mining face, then analyzed the structure of reasonable TML, and identified the optimal position of the TML. Many have attempted to model numerically the reasonable TML locus; based on FLAC numerical software, a method of designing protective coal pillars on a roof roadway based on mining-induced stress boundary line was proposed by Wang et al. [7]. Sun et al. [8] used FLAC3D to simulate the advance abutment stress distribution and the characteristics of coal pillar deformation. The results show that the reasonable position of the TML should be at least 76 m from the main roadway groups. In empirical and in situ monitoring methods, Wei et al. [9] found that the strata structure of the dip pillar area and the asymmetric I-shaped spatial strata structure formed by excavation are primary causes of abnormal stress concentration. The stress distribution in strata overlying the pillar area was obtained by establishing a spatial strata structure model and strike abutment pressure calculation model in the dip pillar area. For the reasonable determination of the TML, i.e., the design of the protective coal pillar between the coalface and the main dip (rise), it is necessary to ensure that the dip (rise) roadways and chamber groups are not within the range of influence of the advance abutment stress. Therefore, the study of the range of influence of advance abutment stress is essential. Meanwhile, the mechanical behaviors in deep coal mining are completely different from the shallow parts, which cause the stress distribution and overlying strata of deep coalfaces to exhibit different characteristics [10–13].In summary, the aforementioned investigations on the determination of the TML are mainly based on experience and lack a theoretical basis, and considering the difference between numerical simulation and site situation, as a result, there are significant discrepancies in the setting design of protective coal pillars on the TML, and it is not conducive to the prevention and control of rock burst risk on site.It is generally known that a rock burst is likely to be induced in case of the total stress (due to superposition of static stress in the coal and dynamic stress induced by mine tremors) exceeding the critical stress [14, 15]. Therefore, according to the method of stress field analysis under seismic wave excitation, the static geological data revealed during roadway excavation were added for the static evaluation. Dynamic evaluation adopted the microseismic monitoring data and stress monitoring systems to monitor dynamic information such as the vibrations and stress signals collected during coalface mining. A combined method of evaluation using both dynamic and static data was proposed for evaluating and optimizing the reasonable TML. To gain an in-depth understanding of mining-induced stress field and rock burst mechanisms around the TML, the following work has been conducted, as reported here: first, the basic information pertaining to a case study of Xingcun Coal Mine (XCM), i.e., geological structures, characteristics of the intrinsic coal properties, and site details of panel 3302, was analyzed. Second, by theoretically analyzing the stress transfer mechanism in the roof structure, a computational model of abutment pressure was established and the calculation method of abutment stress range of influence was proposed. Finally, combined with the results of online stress monitoring, microseismic monitoring, and passive seismic velocity tomography, the reasonable position of TML was determined and optimized to obtain a more realistic evaluation of rock burst risk.


## 2. Stress Distribution around the TML

### 2.1. Structure of the TML

It is generally believed that the determination of TML should ensure that the dip (rise) is in the area of decreasing abutment stress; therefore, the reasonable TML should include the inherent excavation range of influence,l1, of the dip (rise) roadway, the advance range of influence, l2, of the working face, and the minimum safe distance, l, to avoid mutual influence between these two areas (Figure 1).Figure 1
Layout of reasonable terminal mining line.In general, the inherent excavation range of influencel1 of the dip (rise) roadway is fixed. Due to the limitation of geological conditions and monitoring technology, it is difficult to accurately obtain the advance range of influence l2 of the working face. Therefore, the determination of a reasonable value for l2 will be prerequisite to determining the reasonable TML.


### 2.2. Evolution and Load Transfer Mechanism of Overlying Strata above a Single Longwall Panel

The fracture and movement of overlying strata determine the stress distribution and strata behavior. Based on engineering practice and field observation, the overlying strata structure can be divided into three types: OX,F, and T structures [16]. According to whether the key strata is broken or not, the “O” overburden spatial structure can be grouped into two categories: (1) the total OX structure when the primary key strata (PKS) is broken, as shown in Figure 2(a); (2) the half-OX structure when the primary key strata is unbroken, as shown in Figure 2(b). For the total OX structure, the PKS is broken, the load on the KS includes its self-weight, and the load of its controlled strata. Therefore, it can be concluded that 25% of the load was transferred onto the coal seam. For the total OX structure, the PKS is unbroken, it can be considered that 25% of the load induced by the fracture KS and its controlled strata, and 50% of the load induced by the strata above the fracture zone were transferred onto the coal seam.Figure 2
Gob roof structures and stress transfer mechanism after excavation of the longwall panel. (a) Whole OX structure. (b) Half-OX structure.
(a)(b)


### 2.3. Abutment Stress Computation for a Longwall Panel

Based on the load transfer mechanism analysis method before, an abutment stress computational model for the longwall panel is introduced, as shown in Figure3.Figure 3
Abutment stress computational model of the longwall panel.Assuming that theith KS was broken, the load of the ith KS block including its self-weight and its controlled strata is qi, and the load transferred onto the coal seam can be expressed as(1)Qi=qi4=γLiMi+mi4,where γ is the average bulk unit weight of the overlying strata (25 kN/m3), Li denotes the length of the ith KS block, and Mi and mi are the average thickness of the KS and its controlled strata.For brevity, based on engineering experience, it is assumed that the stress transferred onto the coal seam from each KS block is distributed in a shape of an isosceles triangle from the coal face. Thus, the stress transferred onto the coal seam from theith KS block can be written as(2)Qi=σmaxxHcotα,0,Htanα,σmax2−xHcotα,Htanα,2Htanα,0,2Htanα,+∞,where α is the fracture angle, Hd is the HDZ of the goaf area (refer to Zhu et al. [17] for more details), and Δσmax is the greatest stress on the coal seam generated by the goaf area.For the total OX structure, the uppermost KS block bears half the weight of the PKS and its controlled strata, and the total load transferred onto the coal seam by thenth KS block is(3)Qn=qn4=γLnMn+mn+2h1Ln4,where h1 is the average thickness of the PKS.At this time, the stress transferred onto the coal seam from the unbroken PKS is written as(4)QB=qB2=γh1Ln2,where QB is the weight of the PKS in the area of advance of the coalface.The abutment stress on the longwall panel consists of the geostatic stressσg and the stress Δσ transferred from the behind goaf area. Therefore, the abutment stress can be written as(5)σs=σg+Δσ.The geostatic stress on the coal seam can be expressed as(6)σg=γxtanα,0,Htanα,γH,Htanα,+∞.In summary,(1)
For the total OX structure, the abutment stress can be written as(7)σ=σg+∑1nQi+QB.(2)
For the half-OX structure, the abutment stress can be expressed as(8)σ=σg+∑1nQi.


## 2.1. Structure of the TML

It is generally believed that the determination of TML should ensure that the dip (rise) is in the area of decreasing abutment stress; therefore, the reasonable TML should include the inherent excavation range of influence,l1, of the dip (rise) roadway, the advance range of influence, l2, of the working face, and the minimum safe distance, l, to avoid mutual influence between these two areas (Figure 1).Figure 1
Layout of reasonable terminal mining line.In general, the inherent excavation range of influencel1 of the dip (rise) roadway is fixed. Due to the limitation of geological conditions and monitoring technology, it is difficult to accurately obtain the advance range of influence l2 of the working face. Therefore, the determination of a reasonable value for l2 will be prerequisite to determining the reasonable TML.


## 2.2. Evolution and Load Transfer Mechanism of Overlying Strata above a Single Longwall Panel

The fracture and movement of overlying strata determine the stress distribution and strata behavior. Based on engineering practice and field observation, the overlying strata structure can be divided into three types: OX,F, and T structures [16]. According to whether the key strata is broken or not, the “O” overburden spatial structure can be grouped into two categories: (1) the total OX structure when the primary key strata (PKS) is broken, as shown in Figure 2(a); (2) the half-OX structure when the primary key strata is unbroken, as shown in Figure 2(b). For the total OX structure, the PKS is broken, the load on the KS includes its self-weight, and the load of its controlled strata. Therefore, it can be concluded that 25% of the load was transferred onto the coal seam. For the total OX structure, the PKS is unbroken, it can be considered that 25% of the load induced by the fracture KS and its controlled strata, and 50% of the load induced by the strata above the fracture zone were transferred onto the coal seam.Figure 2
Gob roof structures and stress transfer mechanism after excavation of the longwall panel. (a) Whole OX structure. (b) Half-OX structure.
(a)(b)


## 2.3. Abutment Stress Computation for a Longwall Panel

Based on the load transfer mechanism analysis method before, an abutment stress computational model for the longwall panel is introduced, as shown in Figure3.Figure 3
Abutment stress computational model of the longwall panel.Assuming that theith KS was broken, the load of the ith KS block including its self-weight and its controlled strata is qi, and the load transferred onto the coal seam can be expressed as(1)Qi=qi4=γLiMi+mi4,where γ is the average bulk unit weight of the overlying strata (25 kN/m3), Li denotes the length of the ith KS block, and Mi and mi are the average thickness of the KS and its controlled strata.For brevity, based on engineering experience, it is assumed that the stress transferred onto the coal seam from each KS block is distributed in a shape of an isosceles triangle from the coal face. Thus, the stress transferred onto the coal seam from theith KS block can be written as(2)Qi=σmaxxHcotα,0,Htanα,σmax2−xHcotα,Htanα,2Htanα,0,2Htanα,+∞,where α is the fracture angle, Hd is the HDZ of the goaf area (refer to Zhu et al. [17] for more details), and Δσmax is the greatest stress on the coal seam generated by the goaf area.For the total OX structure, the uppermost KS block bears half the weight of the PKS and its controlled strata, and the total load transferred onto the coal seam by thenth KS block is(3)Qn=qn4=γLnMn+mn+2h1Ln4,where h1 is the average thickness of the PKS.At this time, the stress transferred onto the coal seam from the unbroken PKS is written as(4)QB=qB2=γh1Ln2,where QB is the weight of the PKS in the area of advance of the coalface.The abutment stress on the longwall panel consists of the geostatic stressσg and the stress Δσ transferred from the behind goaf area. Therefore, the abutment stress can be written as(5)σs=σg+Δσ.The geostatic stress on the coal seam can be expressed as(6)σg=γxtanα,0,Htanα,γH,Htanα,+∞.In summary,(1)
For the total OX structure, the abutment stress can be written as(7)σ=σg+∑1nQi+QB.(2)
For the half-OX structure, the abutment stress can be expressed as(8)σ=σg+∑1nQi.


## 3. Abutment Stress Computation for Longwall Panel 3302 in Xingcun Coal Mine

### 3.1. Xingcun Coal Mine

The XCM is in Shandong Province, China, and has a length of 6.09 km in the strike direction and a width of 3.45 km in the dip direction. The area of mineable zones is 32.60 km2. The currently mined coal seam lies at a burial depth of 1200 m. The Western No. 3 mining district is distributed along NW-SE direction, which boards fault DF22 (throw varying from 20–105 m) to the east and the western ventilation roadway and track roadway to the west-east; the coal pillar line of railway rise of the Western No. 4 mining district, fault DF16 (throw varying from 0–20 m), and fault DF13 (throw varying from 0–30 m) to the north-west and fault DF14 (throw varying from 5–190 m) to the south-west and south.The Western No. 3 mining district has a length of 3.4–4.2 km in the strike direction and a width of 0.4–3.2 km in the dip direction, and the area of mineable zones is about 3.9 km2. Fault structures are developing in the mining area, which plays a major role in the tectonic characteristics and affects coal mining operations. The fully-mechanized top coal caving method was used to retreat the coalfaces, whose thickness varies from 8.0–8.7 m (8.5 m on average) with an inclination ranging from 5° to 11° (8° on average). Up to 12 Feb. 2016, LW3302, LW3303, LW3307, and LW3308 had been mined out; LW3311 had been developed by the excavation of trackgate and headgate entries, and LW3310 had been retreated by about 570 m (averaging about 1.2 m per day), as shown in Figure 4. In the dip direction, strip coal pillars with a width of more than 100 m were designed between the coalfaces to isolate the goaf and control surface subsidence.Figure 4
Photographs of roadway damage taken after rock burst: (a) excavation face, (b) roof subsidence, (c) coal wall outburst, and (d) belt shift.All such coalfaces were at high risk of rock bursting, making them appropriate sites for case studies for the application of passive seismic velocity tomography.


### 3.2. Intrinsic Burst-Proneness

Intrinsic burst-proneness is an inherent property of coal and can be used to evaluate the risk of a rock burst by assessing the energy accumulation capability of coal. According to the current standard in China (GB/T25217.2 2010), burst-proneness is classified into three levels: high burst-proneness, low burst-proneness, and no burst-proneness. The indices used to evaluate the burst-proneness of coal are the duration of dynamic fracture (DT), the bursting energy index (KE), the elastic strain energy index (WET), and the uniaxial compressive strength (Rc) [18]. Taking coal samples from the No. 3 coal seam in the XCM for testing, the duration of dynamic fracture (DT) was measured to be 47 ms, the bursting energy index (KE) and the elastic strain energy index (WET) were found to be 15 and 5.6, and the uniaxial compressive strength (Rc) was measured at 9.5 MPa, respectively (Table 1). Based on the measured values of these indices, this coal seam has a high burst-proneness.Table 1
Rock burst consequences in Western No. 3 mining district detected byin situ measurement.
DT (ms)KEWETRc (MPa)Results47155.69.5High burst-proneness


### 3.3. Rock Burst Consequences

Table2 shows that there are a total of six rock bursts, two during roadway excavation and four during advances of the coalfaces, causing varying degrees of damage to LW3302, LW3303, and LW3308 in a total of eight areas, such as roof subsidence, damage to single props, coal spalling, and floor heave. Rock bursts 3 (and 6) caused damage to both headgate and coalface simultaneously. Each of the others caused damage to just one area, i.e., the headgate. Hence, rock bursts occurring in the headgate account for most of the events recorded. Figure 5 shows record photographs of the most disastrous rock burst occurring on 26 July 2015. The most disastrous rock burst, occurring on 26 July 2017, caused a 200 m-long damaged section to appear in the headgate and coalface; coal wall movement, floor heave of 0.2–1.5 m, and damage to all single props up to 150 m in front of the coalface ensued. Normally, for moderate or slight rock bursts, lengths of the damaged section of the tailgate lay within 40–90 m during coalface advance. For disastrous rock bursts, the lengths vary from 150–200 m. In general, rock bursts occurred in roadways. While during coalface advancing operation, rock bursts occurred both in roadways and the coalface, with potentially disastrous consequences.Table 2
Burst-proneness determination: No. 3 coal seam in the Xingcun coal mine.
#DateLocation of coalfaceSurrounding geological structureDamage length (m)(a)Extent of damage(b)Description (main damage)12013/1/24Coalface 3303 miningFault SF11140SDamage to 54 single props in the headgate entry22013/1/26Open-off cut excavation of coalface 3308Fault SF104/DF2240SDamage to 17 single props from 14–54 m in the west of the headgate, several anchor and cables broken32013/8/5Roadway excavation 990 m of coalface 3302Fault SF100150DRoadway deformation from 30–180 m behind the excavation face42013/9/8Coalface 3308 miningFault SF10290MTailgate: damage to 60 single props from 6–96 m; roof subsidence 600–800 mm52015/1/30Coalface 3302 mining—60MHeadgate: damage to 43 steel sheds; floor heave 400 mm in front of the headgate62015/7/26Coalface 3302 miningFault with throw smaller than 3 m200D7 people injured, floor heave 0.8–1.5 m at 30 m and floor heave 0.2–0.6 m at a distance of 30–200 m in front of the headgate, and damage to all the single propsColumn 1 assigns a number used to refer to each rock burst in the text.(a)Normal numbers refer to damage to roadways while italic-bold numbers refer to damage to the coalface. For instance, rock burst 12 caused damage to the tailgate (120 m) and coalface (50 m), simultaneously. (b)Disastrous rock burst means that instantaneous deformation (ID) of roadways exceeds 1 m or multiple damage induced by one rock burst, moderate rock burst means that ID ranges from 0.5 m to 1.0 m, and slight rock burst means that ID is less than 0.5 m. D-disastrous, M-moderate, and S-slight.Figure 5
Layout of the MS monitoring system installed in the Western No. 3 mining district. Squares denote temporary stations that can be moved as the coalface advances and triangles represent permanent stations.
(a)(b)(c)(d)On 7 August 2008, an MS monitoring system (SOS, Central Mining Institute (GIG), Poland) was installed in XCM. There was considerable MS activity during coal entry excavation and coalface mining, which makes it ideal for passive tomography imaging. Figure4 illustrates the longwall geometry and geophone locations: there were four geophones (blue circles) used to monitor the seismicity around LW3307 and LW3310, respectively, and these were mounted on the headgate and tailgate roadways allowing them to be moved as the coalface advances.


### 3.4. Site Details of Longwall Panel 3302

Based on the experimental tests of burst-proneness, the average geometrical and geomechanical parameters of the overlying strata over LW 3112 are listed in Table3. Substituting the above parameters into equation (9) [17], the theoretically computed height of the destressed zone can be obtained as 646.11 m:(9)Hd=1+ν1−2νγ2Hh/1−νEh2/3+Hh+H2σc2/E+bσc/b−1,where H is the cover depth, h denotes the thickness of coal seam, E is Young’s modulus, σc is the uniaxial compressive strength of the overlying strata, ν represents Poisson’s ratio, b is the bulking factor, and γ is the average bulk unit weight.Table 3
Average values of geometrical and geomechanical parameters of strata overlying LW 3302.
H (m)h (m)L (m)σc (MPa)γ (kN/m3)υE (GPa)b12008.29584.8250.273.61.5The lithological information above LW 3112 is displayed in Table4. According to the key stratum hypothesis, there are a total of five KS in the overlying strata above LW 3112, as shown in Table 4. To better understand the overlying strata structure and movement, a series of monitoring points were established to observe the surface subsidence. The maximum value of the observed surface subsidence is between 100 mm and 154 mm, which is far less than the maximum surface subsidence in this area. It can be thus inferred that the PKS does not break.Table 4
Lithological information for rock above LW 3302.
LithologyThickness (m)RemarksClay70.19Alluvium60.09Siltstone16.25Sandstone583.54PKSMudstone30.20Siltstone116.79KS 4Fracture strata130.67Mudstone13.46Conglomerate3.68Fine sandstone60.09KS 3Fracture strata33.27Medium sandstone4.49Fine sandstone22.40KS 2Fracture strata42.61Coal seam 20.56Mudstone10.70Medium sandstone27.91KS 1/ main roofMudstone2.80Immediate roofCoal seam 3down8.54Siltstone2.94Immediate floorFinally, considering the theoretically computed HDZ and the key stratum hypothesis, the theoretical HDZ should be corrected to 352.64 m, i.e., the height from the siltstone (above coal seam No. 3) to sandstone (PKS) = 2.8 + 27.91 + 10.7 + 0.56 + 42.61 + 22.4 + 4.49 + 33.27 + 60.09 + 3.68 + 13.46 + 130.67 = 352.64 m.Based on the above analyses and practical geological conditions of LW 3302, all the parameters required for stress calculation are listed as follows:H1 = 16.76 m, L1 = 37.2 m, M1 = 27.91 m, m1 = 53.87 m; H2 = 95.78 m, L2 = 39.4 m, M2 = 22.40 m, m2 = 37.76 m; H3 = 174.79 m, L3 = 66.9 m, M3 = 60.09 m, m3 = 147.81 m; γ = 25.5 kN/m3, h1 = 593.54 m, H = 1200 m. Based on field observation and similar simulation experiments, the fracture angle of overlying strata α is approximately 80°.Substituting the above parameters into equations (1)–(6) and (8), the theoretically computed abutment stress distribution of LW 3302 can be obtained (Figure 6): Figure 6 shows that the peak abutment stress is about 54 MPa and may be found 50 m in front of the coalface. The range of influence of abutment stress is about 255 m, so the TML should be arranged to be at least 255 m in front of the coalface.Figure 6
Theoretically computed abutment stress distribution on LW3302.Figure 7
Layout of pressure sensors installed in the entries of LW3302.


## 3.1. Xingcun Coal Mine

The XCM is in Shandong Province, China, and has a length of 6.09 km in the strike direction and a width of 3.45 km in the dip direction. The area of mineable zones is 32.60 km2. The currently mined coal seam lies at a burial depth of 1200 m. The Western No. 3 mining district is distributed along NW-SE direction, which boards fault DF22 (throw varying from 20–105 m) to the east and the western ventilation roadway and track roadway to the west-east; the coal pillar line of railway rise of the Western No. 4 mining district, fault DF16 (throw varying from 0–20 m), and fault DF13 (throw varying from 0–30 m) to the north-west and fault DF14 (throw varying from 5–190 m) to the south-west and south.The Western No. 3 mining district has a length of 3.4–4.2 km in the strike direction and a width of 0.4–3.2 km in the dip direction, and the area of mineable zones is about 3.9 km2. Fault structures are developing in the mining area, which plays a major role in the tectonic characteristics and affects coal mining operations. The fully-mechanized top coal caving method was used to retreat the coalfaces, whose thickness varies from 8.0–8.7 m (8.5 m on average) with an inclination ranging from 5° to 11° (8° on average). Up to 12 Feb. 2016, LW3302, LW3303, LW3307, and LW3308 had been mined out; LW3311 had been developed by the excavation of trackgate and headgate entries, and LW3310 had been retreated by about 570 m (averaging about 1.2 m per day), as shown in Figure 4. In the dip direction, strip coal pillars with a width of more than 100 m were designed between the coalfaces to isolate the goaf and control surface subsidence.Figure 4
Photographs of roadway damage taken after rock burst: (a) excavation face, (b) roof subsidence, (c) coal wall outburst, and (d) belt shift.All such coalfaces were at high risk of rock bursting, making them appropriate sites for case studies for the application of passive seismic velocity tomography.


## 3.2. Intrinsic Burst-Proneness

Intrinsic burst-proneness is an inherent property of coal and can be used to evaluate the risk of a rock burst by assessing the energy accumulation capability of coal. According to the current standard in China (GB/T25217.2 2010), burst-proneness is classified into three levels: high burst-proneness, low burst-proneness, and no burst-proneness. The indices used to evaluate the burst-proneness of coal are the duration of dynamic fracture (DT), the bursting energy index (KE), the elastic strain energy index (WET), and the uniaxial compressive strength (Rc) [18]. Taking coal samples from the No. 3 coal seam in the XCM for testing, the duration of dynamic fracture (DT) was measured to be 47 ms, the bursting energy index (KE) and the elastic strain energy index (WET) were found to be 15 and 5.6, and the uniaxial compressive strength (Rc) was measured at 9.5 MPa, respectively (Table 1). Based on the measured values of these indices, this coal seam has a high burst-proneness.Table 1
Rock burst consequences in Western No. 3 mining district detected byin situ measurement.
DT (ms)KEWETRc (MPa)Results47155.69.5High burst-proneness


## 3.3. Rock Burst Consequences

Table2 shows that there are a total of six rock bursts, two during roadway excavation and four during advances of the coalfaces, causing varying degrees of damage to LW3302, LW3303, and LW3308 in a total of eight areas, such as roof subsidence, damage to single props, coal spalling, and floor heave. Rock bursts 3 (and 6) caused damage to both headgate and coalface simultaneously. Each of the others caused damage to just one area, i.e., the headgate. Hence, rock bursts occurring in the headgate account for most of the events recorded. Figure 5 shows record photographs of the most disastrous rock burst occurring on 26 July 2015. The most disastrous rock burst, occurring on 26 July 2017, caused a 200 m-long damaged section to appear in the headgate and coalface; coal wall movement, floor heave of 0.2–1.5 m, and damage to all single props up to 150 m in front of the coalface ensued. Normally, for moderate or slight rock bursts, lengths of the damaged section of the tailgate lay within 40–90 m during coalface advance. For disastrous rock bursts, the lengths vary from 150–200 m. In general, rock bursts occurred in roadways. While during coalface advancing operation, rock bursts occurred both in roadways and the coalface, with potentially disastrous consequences.Table 2
Burst-proneness determination: No. 3 coal seam in the Xingcun coal mine.
#DateLocation of coalfaceSurrounding geological structureDamage length (m)(a)Extent of damage(b)Description (main damage)12013/1/24Coalface 3303 miningFault SF11140SDamage to 54 single props in the headgate entry22013/1/26Open-off cut excavation of coalface 3308Fault SF104/DF2240SDamage to 17 single props from 14–54 m in the west of the headgate, several anchor and cables broken32013/8/5Roadway excavation 990 m of coalface 3302Fault SF100150DRoadway deformation from 30–180 m behind the excavation face42013/9/8Coalface 3308 miningFault SF10290MTailgate: damage to 60 single props from 6–96 m; roof subsidence 600–800 mm52015/1/30Coalface 3302 mining—60MHeadgate: damage to 43 steel sheds; floor heave 400 mm in front of the headgate62015/7/26Coalface 3302 miningFault with throw smaller than 3 m200D7 people injured, floor heave 0.8–1.5 m at 30 m and floor heave 0.2–0.6 m at a distance of 30–200 m in front of the headgate, and damage to all the single propsColumn 1 assigns a number used to refer to each rock burst in the text.(a)Normal numbers refer to damage to roadways while italic-bold numbers refer to damage to the coalface. For instance, rock burst 12 caused damage to the tailgate (120 m) and coalface (50 m), simultaneously. (b)Disastrous rock burst means that instantaneous deformation (ID) of roadways exceeds 1 m or multiple damage induced by one rock burst, moderate rock burst means that ID ranges from 0.5 m to 1.0 m, and slight rock burst means that ID is less than 0.5 m. D-disastrous, M-moderate, and S-slight.Figure 5
Layout of the MS monitoring system installed in the Western No. 3 mining district. Squares denote temporary stations that can be moved as the coalface advances and triangles represent permanent stations.
(a)(b)(c)(d)On 7 August 2008, an MS monitoring system (SOS, Central Mining Institute (GIG), Poland) was installed in XCM. There was considerable MS activity during coal entry excavation and coalface mining, which makes it ideal for passive tomography imaging. Figure4 illustrates the longwall geometry and geophone locations: there were four geophones (blue circles) used to monitor the seismicity around LW3307 and LW3310, respectively, and these were mounted on the headgate and tailgate roadways allowing them to be moved as the coalface advances.


## 3.4. Site Details of Longwall Panel 3302

Based on the experimental tests of burst-proneness, the average geometrical and geomechanical parameters of the overlying strata over LW 3112 are listed in Table3. Substituting the above parameters into equation (9) [17], the theoretically computed height of the destressed zone can be obtained as 646.11 m:(9)Hd=1+ν1−2νγ2Hh/1−νEh2/3+Hh+H2σc2/E+bσc/b−1,where H is the cover depth, h denotes the thickness of coal seam, E is Young’s modulus, σc is the uniaxial compressive strength of the overlying strata, ν represents Poisson’s ratio, b is the bulking factor, and γ is the average bulk unit weight.Table 3
Average values of geometrical and geomechanical parameters of strata overlying LW 3302.
H (m)h (m)L (m)σc (MPa)γ (kN/m3)υE (GPa)b12008.29584.8250.273.61.5The lithological information above LW 3112 is displayed in Table4. According to the key stratum hypothesis, there are a total of five KS in the overlying strata above LW 3112, as shown in Table 4. To better understand the overlying strata structure and movement, a series of monitoring points were established to observe the surface subsidence. The maximum value of the observed surface subsidence is between 100 mm and 154 mm, which is far less than the maximum surface subsidence in this area. It can be thus inferred that the PKS does not break.Table 4
Lithological information for rock above LW 3302.
LithologyThickness (m)RemarksClay70.19Alluvium60.09Siltstone16.25Sandstone583.54PKSMudstone30.20Siltstone116.79KS 4Fracture strata130.67Mudstone13.46Conglomerate3.68Fine sandstone60.09KS 3Fracture strata33.27Medium sandstone4.49Fine sandstone22.40KS 2Fracture strata42.61Coal seam 20.56Mudstone10.70Medium sandstone27.91KS 1/ main roofMudstone2.80Immediate roofCoal seam 3down8.54Siltstone2.94Immediate floorFinally, considering the theoretically computed HDZ and the key stratum hypothesis, the theoretical HDZ should be corrected to 352.64 m, i.e., the height from the siltstone (above coal seam No. 3) to sandstone (PKS) = 2.8 + 27.91 + 10.7 + 0.56 + 42.61 + 22.4 + 4.49 + 33.27 + 60.09 + 3.68 + 13.46 + 130.67 = 352.64 m.Based on the above analyses and practical geological conditions of LW 3302, all the parameters required for stress calculation are listed as follows:H1 = 16.76 m, L1 = 37.2 m, M1 = 27.91 m, m1 = 53.87 m; H2 = 95.78 m, L2 = 39.4 m, M2 = 22.40 m, m2 = 37.76 m; H3 = 174.79 m, L3 = 66.9 m, M3 = 60.09 m, m3 = 147.81 m; γ = 25.5 kN/m3, h1 = 593.54 m, H = 1200 m. Based on field observation and similar simulation experiments, the fracture angle of overlying strata α is approximately 80°.Substituting the above parameters into equations (1)–(6) and (8), the theoretically computed abutment stress distribution of LW 3302 can be obtained (Figure 6): Figure 6 shows that the peak abutment stress is about 54 MPa and may be found 50 m in front of the coalface. The range of influence of abutment stress is about 255 m, so the TML should be arranged to be at least 255 m in front of the coalface.Figure 6
Theoretically computed abutment stress distribution on LW3302.Figure 7
Layout of pressure sensors installed in the entries of LW3302.


## 4. Real-Time Stress Monitoring

A real-time stress monitoring system (KJ743) was employed to monitor the vertical stress change during mining of LW 3302 to better understand the actual abutment stress distribution characteristics of the longwall panel.Forty groups of pressure sensors were installed in the headgate and trackgate entry of LW3302, respectively. As shown in Figure7, the space between each of the 40 groups of pressure sensors is 20 m, and each group consists of two pressure sensors installed in 12 m- and 17 m-deep boreholes with a diameter of 45 mm, drilled into the coal seam. As the coalface advanced, the data from typical pressure sensors (20#,21#, and 22#) are displayed in Figure 8, which indicates that most pressures increased significantly as LW3302 advanced to within approximately 200 m to the pressure sensor.Figure 8
Real-time stress monitoring results of the typical pressure sensors20#,21#,and22# as coalface 3302 advances.It can be seen from Figure8 that, taking the 20# stress sensor as an example, the stress change can be divided into three stages with changing coalface advance: ① stable stage (200–160 m from the monitoring point): there is no influence on the stress at the monitoring point by the advance abutment stress of coalface, the value is about 3.3 MPa; ② slowly rising stage (160–40 m from the monitoring point): as the coalface is advanced, the measuring point begins to enter the zone of influence of the advance abutment stress, and the stress increases slowly from 4.4 MPa to 6.3 MPa; ③ sudden increase stage (40–0 m from the monitoring point): the stress shows an overall rising trend with significant change therein, and the value increases from 6.3 MPa to 24.6 MPa. It is worth noting that there is evidence of stress reduction; when the stress reaches the early warning value, large-diameter drilling and blasting were adopted; then, the stress decreased, but would continue to rise soon afterwards. The stress changes at other measuring points are similar to those at pressure sensor 20.


## 5. Microseismic Monitoring

The TML can be determined in combination with the MS monitoring results. According to the theory of rock burst induced by dynamic combined with static loads [19], to prevent the occurrence of a rock burst, we should reduce the static load in the coal and rock mass as much as possible and, at the same time, avoid excessive dynamic load superposition. For the determination of TML, the advance range of influence of mine tremors of each energy level induced by mining is mainly considered. Therefore, the statistical analysis of mine tremor distributions during the mining process of coalface 3302 was conducted (Table 5). Due to the long advance distance of coalface 3302, only statistical results related to mining tremors within 200 m of the TML are listed in Table 5; significant changes in the farthest distance of mine tremors recorded during the mining of coalface 3302 were noted and the minimum, maximum, average, deviation, and standard deviations were 215 m, 260 m, 248 m, 206 m, and 14 m, respectively. Therefore, the normal range of mine tremors is about 262 m, which is determined as the minimum TML.Table 5
Mine tremor distribution during the mining of coalface 3302.
Distance from the open-off cut (m)TimeMaximum distance from the coalface (m)1364–13447.22–7.262501344–13247.17–7.212401324–13047.12–7.162601304–12847.7–7.112501284–12647.2–7.62501224–12446.27–7.12451204–12246.22–6.262151184–12046.17–6.212651164–11846.12–6.162651144–11646.7–6.11235………………Average value248Maximum value260Minimum value215Deviation206Standard deviation14


## 6. Passive Seismic Velocity Tomography

To realize a dynamic evaluation and prediction of rock burst risk of the coalfaces, passive seismic velocity tomography technology was used to detect the mining-induced stress field and its dynamic migration, so as to provide a basis for the distribution of advanced abutment stress [14, 20–23].


### 6.1. Theory of Velocity Tomography

Velocity tomography relies on a simple relationship in which the velocity of a seismic wave is equal to the ray-path length divided by the travel time between the source and the receiver. Therefore, the time is calculated using the integral of the inverse velocity (slowness), from the source to the receiver. Suppose that the ray path of theith seismic wave is Li with a travel time of Ti, the relationship is expressed as [24, 25](10)V=LT⟶VT=L,(11)Ti=∫LidLVx,y,z=∫LiSx,y,zdL,(12)Ti=∑j=1MdijSj,i=1,…,N,where Vx,y,z is the velocity (m/s), Li denotes the ray-path length of the ith seismic wave (m), Ti is the travel time (s), Sx,y,z is the slowness (s/m), dij denotes the distance to the ith ray in the jth grid, N is the total number of rays, and m is the number of voxels.In passive seismic velocity tomography, the source location, and the ray path, can be calculated by using an initial velocity model; however, the velocity, distance, and time in an individual voxel are unknown. Thus, arranging the time, distance, and slowness for each voxel into matrices, the velocity, based on inverse theory, can be calculated by(13)T=DS⟶S=D−1T,where T is the travel time per ray matrix (1 × N), L is the distance per ray per voxel matrix (N × N), and P represents the slowness per grid cell matrix (1 × M).Usually, the inverse problem is either underdetermined (more voxels than rays) or overdetermined (more rays than voxels) [24]. The most effective way to solve this problem is by use of an iterative process; SIRT is an appropriate algorithm for this purpose [25] and was adopted in this research.


### 6.2. Assessment of Rock Burst Hazard

There are two fundamental regions where the occurrence of seismic hazards is frequent in an underground coal mine: one is the highP-wave velocity region, where there is a positive anomaly in the P-wave velocity. Similarly, there is a negative anomaly in the P-wave velocity in areas undergoing stress relief. The other is in obvious velocity gradient regions, where there may be high stress differences.The velocity anomaly is expressed [26] as(14)An=υp−υPaυpa,where υp is the P-wave velocity at a certain point and υPa is the average velocity across the entire model. The relationship between the positive anomalies in the wave velocity, and stress concentration, is shown in Table 6.Table 6
Relationship between positive anomalies of wave velocity and stress concentration.
Burst risk indexStress concentrationPositive velocity anomaly,An (%)Probability of stress concentration0None<5<0.21Weak5–150.2–0.62Moderate15–250.6–1.43Strong>25>1.4The VG value is used to express the extent of variation in the velocity gradient, and the anomaly in the VG value is given by(15)An=VG−VGaVGa,where VGa is the average velocity gradient; similarly, the relationship between the VG anomaly and rock burst hazard is shown in Table 7.Table 7
Relationship between VG value and rock burst hazard.
Rock burst indexStress differenceVG anomaly (%)0None<51Weak5–152Moderate15–253Strong>25


### 6.3. Data Acquisition

Data acquisition for passive velocity tomography was conducted using the MS system, as shown in Figure1, which uses mining-induced seismic events as the energy sources and measures P-wave arrival time after the seismic wave passes through the rock mass. After data reduction, the data were analyzed using the MINESOSTOMO program developed by Gong [26]. To decrease the grid model size and thereby improve the inversion efficiency, the passive seismic velocity tomography on LW 3302 was performed using various stations 2#,3#,7#,8#,9#,and16# and those seismic events located in the target area. Resolution is one of the most important parameters in evaluating the tomograms. Resolution is typically dependent on three factors: the distribution of the velocity, the source-receiver geometry and density, and the wavelength or grid size [24, 25].


### 6.4. Inversion Parameters

For the SOS system, the wavelength is about 26.7 m, the maximum epicenter location error is about 30 m, and the location error in the vertical direction usually exceeds 70 m [20]. In this study, considering that the dip (in the X-direction) and strike (in the Y-direction) lengths of the Western No. 3 Mining District are about 800 m and 1500 m, respectively, and the mining depth (in the Z-direction) of the Western No. 3 Mining District is about 1200 m, a total of 27 × 51 × 3 voxel points, each with a voxel size of 30 m × 30 m × 125 m, was considered. A layer initial velocity model in calculation was assumed in the investigation area, based on geophysical data that has been previously collected, while the constant P-wave velocity equal to 4.1 km/s was assumed to calculate the seismic event location and perturb the first iteration. To reduce indeterminacy, the velocity was restricted to between 2.3 and 6.0 km/s.


### 6.5. Tomography Results

In view of the “7.26” rock burst in 2015, MS events between 26 July and 4 August 2015 were selected for inversion, and plan views of velocity tomograms at the seam level,Z = −1200 m, were generated for about half a month, to evaluate rock burst risk around the TML, as shown in Figure 9.Figure 9
Assessment result through passive seismic velocity tomography: (a) MS events distribution; (b) velocity anomaly inversion result; (c) velocity gradient anomaly inversion result.
(a)(b)(c)It can be seen from Figure9(a) that, after the rock burst, MS events are mainly small mine tremors with a seismic energy of between 102 J and 103 J, and the distribution of MS events is relatively scattered and starts to spread forwards and backwards to the rock burst area, which indicates that the stress and energy are released to a certain extent after the rock burst. The coalface was retreated to a distance of 51 m to the left of the designed terminal line at the end of July. In this area, influenced by high-stress zones formed by the intersection of roadways in the dip of Western No. 3 mining district and advanced abutment stress on the coalface, the coalface suffers high rock burst risk, which means that the stress further transfers to greater depth, causing stress concentration around the TML of coalface 3302, as shown in Figures 9(b) and 9(c). Tomographic velocity images show there are obvious zoned characteristics of velocity anomaly and velocity gradient anomaly: the advance range of influence is about 220 m; considering the inversion accuracy (30 m), the result should be a certain value in the interval from 220 to 250 m.


## 6.1. Theory of Velocity Tomography

Velocity tomography relies on a simple relationship in which the velocity of a seismic wave is equal to the ray-path length divided by the travel time between the source and the receiver. Therefore, the time is calculated using the integral of the inverse velocity (slowness), from the source to the receiver. Suppose that the ray path of theith seismic wave is Li with a travel time of Ti, the relationship is expressed as [24, 25](10)V=LT⟶VT=L,(11)Ti=∫LidLVx,y,z=∫LiSx,y,zdL,(12)Ti=∑j=1MdijSj,i=1,…,N,where Vx,y,z is the velocity (m/s), Li denotes the ray-path length of the ith seismic wave (m), Ti is the travel time (s), Sx,y,z is the slowness (s/m), dij denotes the distance to the ith ray in the jth grid, N is the total number of rays, and m is the number of voxels.In passive seismic velocity tomography, the source location, and the ray path, can be calculated by using an initial velocity model; however, the velocity, distance, and time in an individual voxel are unknown. Thus, arranging the time, distance, and slowness for each voxel into matrices, the velocity, based on inverse theory, can be calculated by(13)T=DS⟶S=D−1T,where T is the travel time per ray matrix (1 × N), L is the distance per ray per voxel matrix (N × N), and P represents the slowness per grid cell matrix (1 × M).Usually, the inverse problem is either underdetermined (more voxels than rays) or overdetermined (more rays than voxels) [24]. The most effective way to solve this problem is by use of an iterative process; SIRT is an appropriate algorithm for this purpose [25] and was adopted in this research.


## 6.2. Assessment of Rock Burst Hazard

There are two fundamental regions where the occurrence of seismic hazards is frequent in an underground coal mine: one is the highP-wave velocity region, where there is a positive anomaly in the P-wave velocity. Similarly, there is a negative anomaly in the P-wave velocity in areas undergoing stress relief. The other is in obvious velocity gradient regions, where there may be high stress differences.The velocity anomaly is expressed [26] as(14)An=υp−υPaυpa,where υp is the P-wave velocity at a certain point and υPa is the average velocity across the entire model. The relationship between the positive anomalies in the wave velocity, and stress concentration, is shown in Table 6.Table 6
Relationship between positive anomalies of wave velocity and stress concentration.
Burst risk indexStress concentrationPositive velocity anomaly,An (%)Probability of stress concentration0None<5<0.21Weak5–150.2–0.62Moderate15–250.6–1.43Strong>25>1.4The VG value is used to express the extent of variation in the velocity gradient, and the anomaly in the VG value is given by(15)An=VG−VGaVGa,where VGa is the average velocity gradient; similarly, the relationship between the VG anomaly and rock burst hazard is shown in Table 7.Table 7
Relationship between VG value and rock burst hazard.
Rock burst indexStress differenceVG anomaly (%)0None<51Weak5–152Moderate15–253Strong>25


## 6.3. Data Acquisition

Data acquisition for passive velocity tomography was conducted using the MS system, as shown in Figure1, which uses mining-induced seismic events as the energy sources and measures P-wave arrival time after the seismic wave passes through the rock mass. After data reduction, the data were analyzed using the MINESOSTOMO program developed by Gong [26]. To decrease the grid model size and thereby improve the inversion efficiency, the passive seismic velocity tomography on LW 3302 was performed using various stations 2#,3#,7#,8#,9#,and16# and those seismic events located in the target area. Resolution is one of the most important parameters in evaluating the tomograms. Resolution is typically dependent on three factors: the distribution of the velocity, the source-receiver geometry and density, and the wavelength or grid size [24, 25].


## 6.4. Inversion Parameters

For the SOS system, the wavelength is about 26.7 m, the maximum epicenter location error is about 30 m, and the location error in the vertical direction usually exceeds 70 m [20]. In this study, considering that the dip (in the X-direction) and strike (in the Y-direction) lengths of the Western No. 3 Mining District are about 800 m and 1500 m, respectively, and the mining depth (in the Z-direction) of the Western No. 3 Mining District is about 1200 m, a total of 27 × 51 × 3 voxel points, each with a voxel size of 30 m × 30 m × 125 m, was considered. A layer initial velocity model in calculation was assumed in the investigation area, based on geophysical data that has been previously collected, while the constant P-wave velocity equal to 4.1 km/s was assumed to calculate the seismic event location and perturb the first iteration. To reduce indeterminacy, the velocity was restricted to between 2.3 and 6.0 km/s.


## 6.5. Tomography Results

In view of the “7.26” rock burst in 2015, MS events between 26 July and 4 August 2015 were selected for inversion, and plan views of velocity tomograms at the seam level,Z = −1200 m, were generated for about half a month, to evaluate rock burst risk around the TML, as shown in Figure 9.Figure 9
Assessment result through passive seismic velocity tomography: (a) MS events distribution; (b) velocity anomaly inversion result; (c) velocity gradient anomaly inversion result.
(a)(b)(c)It can be seen from Figure9(a) that, after the rock burst, MS events are mainly small mine tremors with a seismic energy of between 102 J and 103 J, and the distribution of MS events is relatively scattered and starts to spread forwards and backwards to the rock burst area, which indicates that the stress and energy are released to a certain extent after the rock burst. The coalface was retreated to a distance of 51 m to the left of the designed terminal line at the end of July. In this area, influenced by high-stress zones formed by the intersection of roadways in the dip of Western No. 3 mining district and advanced abutment stress on the coalface, the coalface suffers high rock burst risk, which means that the stress further transfers to greater depth, causing stress concentration around the TML of coalface 3302, as shown in Figures 9(b) and 9(c). Tomographic velocity images show there are obvious zoned characteristics of velocity anomaly and velocity gradient anomaly: the advance range of influence is about 220 m; considering the inversion accuracy (30 m), the result should be a certain value in the interval from 220 to 250 m.


## 7. Reasonable Determination of the TML

In summary, combined with the results of theoretical analysis (255 m), stress monitoring (200 m), MS monitoring (262 m), and passive seismic velocity tomography (220 to 250 m), it can be finally determined that the width of the protective coal pillar in the TML of coalface 3302 should be at least 262 m; therefore, based on above analysis, a flow chart for the reasonable determination of the TML is provided based on the stress field under seismic wave excitation, as illustrated in Figure10.Figure 10
Flow chart of reasonable determination of terminal mining line based on the stress field with the seismic wave.


## 8. Conclusion

With widespread mine-out, the TML suffers frequent rock burst disasters, which has become a typical kind of rock burst encountered in deep coal mining in China. Investigation of the advance abutment stress distribution can provide a theoretical basis for mine design based on rock burst prevention, thus greatly enhancing mining safety in deep coal mines.According to whether the key strata are broken or not, the overburden spatial structure can be grouped into two categories: the total OX structure and the half-OX structure. The overburden spatial structure determines the load transfer mechanism, the transferred stress increment from the behind goaf area is considered, and an abutment stress computational model for the longwall panel is proposed in the present work. Using this model, the abutment stress on LW3302 in Xingcun Coal Mine was determined to be approximately 50 MPa (peak stress) at a point some 50 m in front of the coalface with a range of influence of 255 m.According to the technique of stress field measurement under seismic wave monitoring, a dynamic evaluation for the determination of the TML was made. With both the static and dynamic information from the excavation, the proposed evaluation method overcomes the limitation of traditional static evaluation and provides a more accurate basis for rock burst control during mining activities.Combined with the results of theoretical analysis (255 m), online stress monitoring (200 m), microseismic monitoring (262 m), and passive seismic velocity tomography (220 to 250 m), it can be finally determined that the width of the protective coal pillar for the TML of coalface 3302 should be at least 262 m. The method for dynamic assessment of the most reasonable TML developed in this study can be used as a basis for further research into appropriate protective coal pillar design for use in deep coalfaces of rock-burst-prone coal mines.



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*Source: 3929004-2021-08-26.xml*