Coupled mechanism of compression and prying-induced rock burst in steeply inclined coal seams and principles for its prevention

Tunnelling and Underground Space Technology 98 (2020) 103327

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Coupled mechanism of compression and prying-induced rock burst in steeply inclined coal seams and principles for its prevention Shengquan Hea,b,c, Dazhao Songa,b, Liming Qiua,b

⁎,1

T

, Xueqiu Hea,b,d, Jianqiang Chene, Ting Rend, Zhenlei Lia,b,

a

School of Civil and Resources Engineering, University of Science & Technology Beijing, Beijing 100083, China Key Laboratory of Ministry of Education for Efficient Mining and Safety of Metal Mine, University of Science & Technology Beijing, Beijing 100083, China c Department of Mining and Materials Engineering, McGill University, Montreal H3A 0E8, Canada d School of Civil, Mining and Environmental Engineering, University of Wollongong, Wollongong, NSW 2522, Australia e Shenhua Xinjiang Energy Company Limited, Urumqi 830027, China b

A R T I C LE I N FO

A B S T R A C T

Keywords: Rock burst Steeply inclined coal seam Elastic deformation energy Dynamic instability Stress evolution Prevention strategy

Rock bursts frequently occur in steeply inclined and extremely thick coal seams (SIETCS), posing severe challenges to safe mining. To reduce the risk of rock burst in SIETCS, this study investigated the mechanisms of the rock bursts occurrence in SIETCS and formulated the principles for its prevention. To this end, field investigation, geophysical monitoring, theoretical analyses, and numerical simulation were employed. Mechanical models have been developed for a “steeply inclined suspended roof structure” and a “steeply inclined suspended rock pillar structure”, which are relevant to the rock burst mechanisms. The elastic deformation energy distribution functions for both models have been obtained, and the factors influencing the elastic deformation energy have been analyzed. The sources of microseismic (MS) events associated with rock bursts monitored in the typical SIETCS (with a dip angle of 87°) are mainly concentrated in the roof and interlayer rock pillar, making up 17.0% and 60% of the events recorded, respectively. The elastic deformation energy of the roof and rock pillar is mainly influenced by the dip angle of the coal seam, the lateral pressure coefficient, and the supporting force coefficient. The peak stress of the coal body at the compressive and prying area is 1.7 times of the horizontal tectonic stress. The minimum normal and tangential dynamic load stresses generated by the recorded rock bursts are 84.5 MPa and 48.6 MPa, respectively; such stress levels exert strong destructive forces when superimposed with static stress. The analytical results of the failure law of rock burst, MS monitoring, mechanical model, numerical simulation, and elastic deformation energy function of the typical SIETCS identify the main causes of rock burst as the high static stress of a coal body under the coupled action of compressive and prying effects of roof and rock pillar and the dynamic stress caused by breakage of the roof and rock pillar. The damage models and the damage process by which a rock burst is induced have been constructed. Three mechanisms by which a rock burst can occur in SIETCS are proposed. Finally, prevention principles of load-reduction and prying-reduction for rock burst in SIETCS have been developed.

1. Introduction Rock burst is one of the main dynamic disasters in underground engineering, such as coal mines, hard-rock mines, transportation tunnels, and hydro-power caverns (Cai et al., 2018; Simser, 2019; Feng et al., 2019; Pu et al., 2019; Xue et al., 2020). A rock burst in hard rock mines, transportation tunnels, and hydro-power caverns refers to smallscale dynamic rock damage and ejection that suddenly occur around excavations, which is most common occurrences because hard rock can

store more strain energy (Blake and Hedley, 2009; Zhu et al., 2016; Liu et al., 2019). In some cases, may cause damage to the equipment and personal injury. Rock burst in coal mines refers to the sudden and severe structural damage to the coal seam and the roadway that releases excessive elastic strain energy and expels large amounts of coal and rock into a roadway or working face (Wang et al., 2013). The scale and severity of the damage usually is greater than hard rock engineering, a strong rock burst can even destroy more than a kilometer roadway with serious casualties and equipment damage (Zhu et al., 2016).

⁎

Corresponding author. E-mail address: [email protected] (D. Song). 1 Address: University of Science and Technology Beijing, No.30, Xueyuan Road, Haidian District, Beijing, China. https://doi.org/10.1016/j.tust.2020.103327 Received 22 June 2019; Received in revised form 26 January 2020; Accepted 28 January 2020 Available online 21 February 2020 0886-7798/ © 2020 Elsevier Ltd. All rights reserved.

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approved the mining of steeply inclined coal seams with complex coal and rock conditions. Steeply inclined coal seams occur widely in many coal-producing areas, including Xinjiang, Ningxia, and Gansu in China, Asturias in Spain, and Lorraine in France (He et al., 2019; Deng and Wang, 2014; Díez and Álvarez, 2000; Driad-Lebeau et al., 2005; Heib, 2012). At steeply inclined and extremely thick coal seams (SIETCS), rock burst is a major problem. Because of their special characteristics of high horizontal geostress and complex geological conditions, the critical depth for rock burst is less than that of most coal mines. However, the above-mentioned studies mainly focused on horizontal and gently inclined coal seams, although several important research results have been achieved for steeply inclined coal seams. Furthermore, most of these studies focused on a single stope or a single factor of multiple stopes. Few systematic studies investigated the mechanism of rock burst when adjacent seams are simultaneously mining deep under the condition of goaf filling, especially in SIETCS. Because of the characteristic sedimentary structure and stress distribution of SIETCS, a suspended roof and interlayer rock pillar will form when two coal seams are mined together. The asymmetric distribution of the surrounding rock results in asymmetric roof and rock pillar deformation, failure, energy distribution, and stress transfer. This asymmetric action can easily cause local deformation of the surrounding rock, which can lead to dynamic instability of the roof and floor. The research results of our team identified significant differences between the precursor characteristics of a rock burst in SIETCS and those of a rock burst in horizontal/gently inclined seams (He et al., 2019). Field practice indicates that if a rock burst is assumed to occur by the same mechanism in SIETCS (when two coal seams are mined together under gob filling conditions) often results in a waste of manpower and material resources as it cannot achieve effective prevention. The mechanisms of rock burst have not yet been investigated in steeply inclined coal mines with such coal and rock occurrences and mining conditions. Therefore, it is of theoretical and practical value to systematically investigate the mechanism that causes rock burst, and the principles underlying rock burst prevention in such coal seams. This study investigated the typical SIETCS of No. B1 + 2 and No. B3 + 6 coal seams (with a dip angle of 87°) in the Wudong coal mine (WCM), China, as engineering background. The mechanisms of the rock burst under the condition of two coal seams mined together and goaf filling have been systematically analyzed. This analysis was based on an analysis of the damage characteristics of the rock burst, spatial distribution of microseismic (MS) sources, mechanical model, and elastic deformation energy distribution function for the suspended roof and interlayer rock pillar, numerical simulation, and dynamic stress. Then, the damage models and the damage process, in which the rock burst can be induced, have been developed, and a load-reduction and pryingreduction method for the rock burst prevention and control have also been developed based on the mechanisms of the rock burst. The parameters of acoustic emission (AE) energy and the number of AE pulses per million tons of coal production, daily total MS energy, number of MS events per meter and roadway deformation conditions have been used to test the pressure relief effect of the resulting load-reduction and prying-reduction project. This study provides a reference for rock burst prevention and control in other SIETCS with similar geological and mining conditions.

Simultaneously, rock bursts in coal mines can induce a series of secondary disasters, such as coal and gas outbursts, and gas explosions. For example, on February 14, 2005, a gas explosion induced by a rock burst occurred in Sunjiawan coal mine, killed 214 people, injured 30 people (Li et al., 2016). Therefore, the research of rock burst in coal mine is of great significance. A variety of geological factors, such as strong main roof and floor strata, burst tendency of coal rock, depth of cover, faults, fold, proximity of important tectonic structures, sandstone channels, seam rolls, and major joint sets, as well as mining factors, such as mine design parameters and mining retreat rate, contribute to rock burst (Jiang et al., 2014; Konicek and Schreiber, 2018; Suorineni et al., 2014; Wang et al., 2019a; Zhang et al., 2017; Jiang et al., 2019). Many studies have explored the mechanisms underlying rock burst (Cook, 1963; Singh, 1988; Neyman et al., 1972; Petukhov and Linkov, 1979; Salamon and Wagner, 1979; Zhang et al., 1991; Dou et al., 2014; Lu et al., 2019), and have identified three major groups, which contribute to the rock burst mechanisms, namely energy release, stress concentration, and seismic activity. In the energy release mechanism, Zhang et al. (2017) and Huang et al. (2018) reported that the cantilever beam structure formed over the goaf not only concentrates a large amount of vertical and lateral abutment pressure (static pressure) on the working face but can also store considerable elastic deformation energy within massive strata, which can initiate coal burst events. Huang et al. (2017) found that when the stored elastic energy of the hard roof reached a maximum and was abruptly released, dynamic disasters such as rock burst occurred. Zhang et al. (2017) concluded that the coal burst was caused by a dynamic and unstable release of energy within the overstressed rock mass/coal during the mining process. In the stress concentration mechanism, rock burst occurrences are affected by how the overstress is formed and result from a combination of static and dynamic stress concentrations (Dou et al., 2014; He et al., 2017; Kong et al., 2019; Li et al., 2018a). Rock bursts generally occur during periods of strong dynamic stress disturbance, such as hard-roof rupture, fault slide, pillar failure, working-face mining, or when roadways are being advanced (Li et al., 2014; He et al., 2012; He et al., 2017). A hard and thick roof stratum exerts a major influence over coal breakage since it imposes a high dynamic stress load on the coal mass when the roof fractures (Feng et al., 2011; Wang et al., 2019c). Marcak (2012) reported that in many underground mines in Poland, the stress of the curved roof was a dangerous source of rock burst. Lai et al. (2016) concluded that the stress concentration caused by the bending of rock pillars between steeply inclined seams was the main reason for frequent dynamic disasters. In the seismic activity mechanism, Dou and He (2001) observed that coal bursts occurred hours after the initiating seismic events. Chen et al. (2012) reported that the violent shock produced by hard-roof rupture or slip instability, during which a large amount of elastic energy was suddenly released, easily induced rock burst. Wang et al. (2019b) stated that seismic events contributed to a large proportion of rock bursts. The current control techniques for the prevention and control of rock burst can be classified into two groups: preventative controls, such as mine layout design, pillar design, and protective seams in multiple seam mining, and mitigating controls, such as destress blasting, large diameter drilling, hydraulic fracturing, and water infusion. These control measures aim to achieve one or more of the following: avoid high static stress concentration, reduce the magnitude of dynamic events induced primarily by strata breakage, release strain energy stored in coal and rock masses, move the high stress concentration further away from the working face and the wall of the roadway, provide artificial discontinuities to enable cracks to form on the roof, and weaken the strength of the surrounding rock (Wei et al., 2018; Konicek et al., 2011; Ju and Li, 2008; Zhang, 2010; Calleja and Porter, 2016; Chen et al., 2019). With the reduction of horizontal and gently inclined coal seam reserves and the growing demand for coal in China, the government has

2. Engineering background and damage characteristics 2.1. Engineering overview The geological structure of the WCM in China is shown in Fig. 1. Fig. 1 shows the syncline structure, and its south mining area is a typical SIETCS group. The coal seam used in this study is No. B3 + 6, with a dip angle of 87°. The results of in-situ stress measurements are shown in Table 1. The maximum horizontal principal stress σH at each measuring point exceeds the vertical stress σV, and the maximum 2

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No.B3+6

Synclinal axis No.B1+2

Fig. 1. Geological structure representation of the Wudong coal mine (WCM), China.

horizontal principal stress σH was about 1.74–1.90 times of the vertical stress σV, showing that a horizontal tectonic stress field is in operation. Coal seams No. B3 + 6 and No. B1 + 2 were mined alternately. Coal seam No. B3 + 6 was mined first. The mining height and length were 25 m and 2500 m, respectively. The goaf above the mining level is filled with loess (please refer to He et al. 2019 for more details). According to the record, coal seam No. B3 + 6 experienced a total of five rock bursts.

2.3. Analysis of the characteristics of rock burst damage Table 2 shows the statistical results of the rock burst damage in coal seam No. B3 + 6. These data show that the critical mining depth for rock burst occurrence in coal seam No. B3 + 6 is only 300 m, which is less than the depth rock bursts occur in most mines. The rock bursts exhibited obvious directionality; the rock burst “2⋅1” showed directionality from north to south of the roadway, and from the top to the bottom. The high-intensity rock burst caused the belt and tail of the belt conveyor to move southward, the two walls to converge, and a much larger scope and degree of damage to the roadway on the roof side than on the rock-pillar side. Based on the combination of the damage seen at the roadway after the occurrence of the rock burst and the data from MS monitoring, it can be concluded that the main factor that may cause this phenomenon is the large “load”. The “load” of the horizontal tectonic stress and the overburden was transmitted to the coal body through the hard roof. As such, the force and energy were transmitted along the coal body in the upper part of the roadway, and the coal body was compressed, thus causing the north shoulder of the roadway and the roof to sink (Fig. 4(c)). The characteristics of the damage in other rock bursts generally included converging of the two walls, heaving of the floor at the bottom corner of the south side of the roof roadway, bending of the hydraulic prop to the north, and the U-shaped shed shrinking, showing that the impact mainly originated from the south of the roadway and pointed toward the north (Fig. 4(a), (b), and (d)). The main reasons for the preceding described phenomena may be because the interlayer hard-rock pillar is affected by the mining disturbance, and the occurrence of the instability and bending deformation, resulting in a prying effect on the coal body, thus causing stress

2.2. Microseismic monitoring The ARAMIS M/E monitoring system was used to monitor MS events occurring at the WCM, and the data can be used to effectively identify the areas at risk of rock burst (Xu et al., 2017). The installation positions of the probes within the MS monitoring system when rock burst “4⋅26” (rock burst designators refer to the month and day of occurrence) occurred is shown in Fig. 2. When rock burst “2⋅27” occurred, the MS monitoring system was not installed. Therefore, this rock burst is not described in the following source description. Fig. 3(a) shows that the sources that induced the rock bursts are located on the hard roof and hard-rock pillar. Vertically, the source of the rock burst is located in the corresponding area of the coal seam in the mining stage, i.e., the area subjected to a large load and affected by prying. The plan view (Fig. 3(b)) indicates that the MS events and the rock bursts are mainly concentrated on the hard roof and hard-rock pillar. Statistical analysis of the MS events one month before each rock burst indicated that the proportions of MS sources on the roof, coal seam, and rock pillar were 17.0%, 23%, and 60%, respectively.

Table 1 Measured tectonic stresses. Measure point

1# 2#

Level

+450 +450

σH

σh

σV

Stress value/MPa

Direction/°

Angle/°

Stress value/MPa

Direction/°

Angle/°

Stress value/MPa

Direction/°

Angle/°

15.77 15.43

158.51 160.50

15.92 13.59

10.21 11.27

76.01 76.27

15.01 8.63

9.27 9.49

192.99 178.77

69.43 65.03

Note: σH represents the maximum horizontal principal stress, σh represents the minimum horizontal principal stress, and σV represents the vertical stress. 3

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MS sensor

11

9

1 2

10 No. B3+6

3

Roof Rock pillar

200m

12

No. B1+2

4 7

5 68

Sensors are arranged at the + 450 horizontal working face Sensors are arranged at the other horizontal working face Fig. 2. Sketch of the distribution of microseismic (MS) monitoring probes in the WCM until 26 April 2017.

model of the “steeply inclined suspended roof structure”, as shown in Fig. 8. A SIETCS is mined from the surface to deeper depths. Therefore, a coordinate system has been developed (Fig. 8), where the end point of the surface roof is the origin, and x is the distance from the surface downward along the axis of the roof. The left side of the roof is affected by the horizontal tectonic stress Fd and the gravitational load of the overlying strata G, which can be simplified as the load F1(x) perpendicular to the inclined direction of the coal seam. According to the study of in situ stress in SIETCS, the horizontal tectonic stress Fd is A times the load of the overlying strata G, Fd = Aρc gx sin θ . The load on the roof along the X-axis passes through the roof section perpendicular to the Y-axis, and the bending moment is zero. Therefore, the axial load was not considered. Assuming that the width of the roof at any section is the unit length, the load on the top of the roof at any section can be expressed as follows:

concentration and energy accumulation at corresponding positions on the rock pillar and the coal body. When the strength limit or the energy storage limit of the coal and rock mass were reached, the hard-rock pillar failed at the next external disturbance, and rock bursts were induced (He et al., 2019). The damage range of each rock burst as shown in Fig. 5, the rock burst damage started both at a certain distance from the working face, and at the working face. The area damaged by the rock bursts was mostly concentrated at 0–209 m in front of the working face. This result indicates that this area is easily affected by mining disturbance and the stress in this area is higher than elsewhere. Therefore, in the process of rock burst prevention and control, measures should be taken to enhance the pressure relief and support of the coal body, hard roof and hardrock pillar within the scope of 209 m in front of the working face. Such measures should reduce the load and the prying effect imposed on the coal body, thus reducing the degree of stress concentration and the risk of rock burst. Simultaneously, through statistical analysis of the damage range of coal seams with different dip angle, it is concluded that when mining with the same method, the damage range in SIETCS is larger than that in typical horizontal and gently inclined coal seams in eastern China, as shown in Fig. 6 (Li et al., 2018b).

F1 (x ) = ρc gx sin θ (A sin θ + cos θ) (0 < x < L + x 0)

(1)

where ρc represents the average density of the overlying strata, g represents gravitational acceleration, and θ represents the dip angle of the coal seam. The supporting force σy, provided by the plastic zone of the coal body to the roof, can be calculated using an expression, which incorporates the stress in the plastic zone and its width (Hou and Ma, 1989). The roof-supporting force σy(x) in the plastic zone (along the Yaxis) can be expressed as follows:

3. Mechanical analysis of the hard roof and hard-rock pillar The damage characteristics of the rock bursts and the MS monitoring results in Sections 2.2 and 2.3 indicate that the rock bursts in SIETCS are closely related to the suspended roof and rock pillar. The degree of stress concentration in the corresponding area of roof and rock pillar was higher. Therefore, it is necessary to analyze the stability of the suspended roof and rock pillar.

2 tan φ c P c σy (x ) = ⎜⎛ + X ⎟⎞ e λm (x − L) − λ⎠ tan φ ⎝ tan φ

(2)

where λ represents the lateral pressure coefficient, which is the ratio of vertical stress to horizontal stress, m represents the thickness of the coal seam, c represents the cohesion at the coal-rock interface, φ represents the angle of internal friction at the coal-rock interface, L represents the length of the roof, and PX represents the supporting force of the loess on the coal. In reality, the degree of compaction of the loess differs at different positions. Therefore, PX can be expressed as follows:

3.1. Mechanical model and elastic deformation energy distribution function of the hard roof After mining, the goaf was filled with loess with different degrees of compaction from top to bottom, resulting in asymmetrical stress distribution in the suspended roof. In this study, a physical model of the “steeply inclined suspended roof structure” (Fig. 7) has been developed, which considers the high horizontal tectonic stress and the weight of the overlying strata. Due to the lower part of the suspended roof extends from the working surface to the coal body. Therefore, it can be regarded as fixed. The plastic zone of the coal body in the fixed area exerts a supporting force to the roof. To facilitate the investigation of the influence of the above-mentioned forces on the roof, this study did not consider the self-weight of the roof and has developed a mechanical

PX = kρ1 gLsin2 θ

(3)

where ρ1 represents the density of the loess, and k represents a supporting force coefficient between 0 and 1. The larger the value of k, the more compact the filling will be. Therefore, k = 1 indicates that the supporting force provided to the roof by the compacted loess reaches a maximum, whereas at k = 0, the loess fill provides zero support to the roof. The width of the plastic zone in the coal body, x0, is: 4

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Fig. 3. (a) Elevation view of rock burst distribution, and (b, c) plan view of MS events and rock burst distribution. c

x0 =

⎛ nF1 (L) + tan φ ⎞ λm In ⎜ ⎟ c P 2 tan φ ⎜ + λX ⎟ ⎠ ⎝ tan φ

σy (L) = F2 (L) = (4)

PX λ

(5)

Loess filling ranges from x = 0 to x = L. When x = 0, the supporting force of the loess to the roof is considered zero. Therefore, to simplify the problem, the supporting force of the loess to the roof was considered to increase linearly along the X-axis. Hence, the supporting load F2(x) is:

where I represents the moment of inertia, and n represents the stress concentration factor. The coal-loess interface is at × = L, at which point, it is considered that the supporting force of the coal body to the roof equals that of the loess to the roof:

F2 (x ) = 5

PX x λL

(6)

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Fig. 3. (continued)

The mechanical model of the “steeply inclined suspended roof structure” described above can be simplified as shown in Fig. 9. In this model, the roof is simplified as a cantilever beam, which is subjected to the loads F1(x), F2(x), and σy(x) at different locations. The load distribution causes bending moment M(x) at location x along the beam, which can be described by the following equation: 3

the suspended rock pillar. To facilitate force analysis of the rock pillar, a mechanical model of a “steeply inclined suspended rock pillar structure” (Fig. 10) has been developed, which comprehensively considers the stresses imposed on it. The stress on the surface of the rock pillar in the goaf includes the residual stress after the attenuation of horizontal tectonic stress and the pressure produced by the loess on the rock pillar, which can be simplified as the load F3(x) perpendicular to the rock pillar. Assuming that the width of the rock pillar at any section is the unit length, according to Newton's third theorem, it can be derived that the load on the upper surface of the rock pillar at any section is:

3

ρc gx sin θ (A sin θ + cos θ ) P x ⎧ − 6XλL 6 ⎪ (0 ⩽ x ⩽ L) ⎪ ⎪ 2 PX L (x − L ) 2 3 sin θ (A sin θ + cos θ ) ρ gx λm c ⎪ c 3 − − 2 tan φ + ⎪ 6 2λ tan φ M (x ) = − 2 tan φ ( x L ) ⎨ − 1⎞ ⎪ ⎛e λm ⎠ ⎪ ⎝ λm c P c (x − L)2 ⎪ + 2 tan φ tan φ + λX (x − L) + 2 tan φ ⎪ ⎪ (L < x ⩽ L + x 0) ⎩

(

(

)(

PX λ

)

F3 (x ) =

)

(7)

F4 (x ) =

(

(

)

)(

)

kρ gx sin θ PX x = 1 (0 < x < L) λL λ

(10)

2 tan φ c P c σy2 (x ) = ⎜⎛ + X ⎟⎞ e mλ (x − L) − + ρ2 g (x − L) sin θ cos θ λ⎠ tan φ ⎝ tan φ

2 PX λ

(9)

Similar to the method for calculating the roof-supporting force in the plastic zone, the expressions for the supporting forces σy2(x) and σy3(x) given by the plastic zone to the left and right sides of the rock pillar are:

2

3

(0 < x < L)

The force at any section on the right side of the rock pillar can be expressed as follows:

According to the relationship between bending moment and elastic deformation energy, the elastic deformation energy at any x along the roof can be expressed as follows: ρ gx sin θ (A sin θ + cos θ ) P x3 1 ⎧ ⎡ c − 6XλL ⎤ 2EI 6 ⎪ ⎣ ⎦ ⎪ (0 ⩽ x ⩽ L) ⎪ 2 ⎪ PX L (x − L ) 2 ⎡ ρc gx 3 sin θ (A sin θ + cos θ) λm c 3 ⎪ − − 2 tan φ + ⎪ ⎢ 6 2λ tan φ U (x ) = ⎨ 1 ⎢ 2 tan φ (x − L) − 1⎞ ⎪ 2EI ⎢ ⎛e λm ⎠ ⎪ ⎢ ⎝ ⎢ ⎪ λm c P c (x − L)2 ⎢ + + λX (x − L) + 2 tan φ ⎪ 2 tan φ tan φ ⎣ ⎪ ⎪ (L < x ⩽ L + x 0) ⎩

kρ1 gx sin2 θ + ρ1 gx sin θ cos θ λ

⎤ ⎥ ⎥ ⎥ ⎥ ⎥ ⎥ ⎦

(11)

PX ⎞ 2 tan φ (x − L) ⎟ e m2 λ

c σy3 (x ) = ⎜⎛ + tan φ λ⎠ ⎝

−

c tan φ

(12)

where ρ2 represents the density of the coal, and m2 represents the thickness of the coal seam. The mechanical model of the “steeply inclined suspended rock pillar structure” described above is a simplified model, and is shown in Fig. 11. The rock pillar is simplified as a cantilever beam, which is subjected to the loads F3(x), F4(x), σy2(x), and σy3(x) at different locations. The load distribution causes the bending moment M(x) at location × along the beam, which can be described by the following equation:

(8) where E represents the elastic modulus. 3.2. Mechanical model and elastic deformation energy distribution function of the hard-rock pillar The goaf was filled with loess at different degrees of compaction from top to bottom, resulting in an asymmetrical stress distribution on 6

+450

+450

2/1/2017

4/26/2017

+475

3/13/2015

+450

+500

2/27/2013

11/24/2016

Mining level (m)

Date

7

1568

1824

2026

1995

1959

Location of working face (m)

+431

+460

+460

Rock pillar

Roof

Rock pillar

Rock pillar

–

–

+480

Location of source

Elevationof seismic source (m)

2.2 × 106

Tailentry Headentry

2.1 × 108

Tailentry

Headentry

Tailentry

Headentry

9.5 × 106

1790–1808 (18) 1500–1520 (20) 1510–1565 (55)

1750–1959 (2 0 9) – 1920–1995 (75) 1840–1950 (110) 1980–2025 (45) – 1630–1824 (194)

Tailentry Headentry Tailentry

1750–1959 (2 0 9)

Damage range (m)

Headentry

Roadway

5 × 108

–

MS energy (J)

Table 2 Consequences of rock burst in coal seam No. B3 + 6 as detected by in-situ measurements.

3

48

16

0 0

1

– 0

0

0

Distance from damage location to working face (m)

– Roof moved and 2 × 2 × 1 m string bag structure appeared, lower wall moved 0.5–1.3 m, upper wall moved 0.3–0.7 m, belt conveyor moved 0.3–1.2 m southward, roof subsided 0.7–1 m Lower wall moved 0.3–0.5 m, floor heaved 0.2–0.3 m, damage to hydraulic props Dislocation of “H” frame of belt conveyor, reversed loader and slideway changed direction Floor heaved 0.3 m, closure of relief groove of south side of tail entry, upper wall moved 0.4 m, damage to hydraulic prop, damage to relief valves of hydraulic support

Roof subsided 0.3–0.5 m, floor heaved 0.2–0.6 m, upper wall moved 0.2–0.5 m, 3 × 1 × 1.5 m cavity appeared in upper wall, reversed loader moved 0.3 m northward Roof subsided 0.1–0.3 m, lower wall moved 0.2 m, 13 I-beams severely deformed – Upper-wall shoulder subsided 0.3 m, U-shaped shed shrank 0.2 m, damage to hydraulic props Upper wall moved 0.3 m, lower wall moved 0.5 m, floor heaved 0.4 m, damage to hydraulic props, top beam weld loosened, direction change of reversed loader

Form and extent of damage

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a

3

(ρ1 + ρc ) gx sin θ cos θ ⎧ 6 ⎪ (x ⩽ L) ⎪ ⎪ ρ gL2 sin θ cos θ (x − 2 L ) 1 ρ gx 3 sin θ cos θ λm c P 3 ⎪ + c + ( 2 tan φ )2 ( tan φ + λX ) 2 6 ⎪ ⎪ 2 tan φ (x − L) ⎪ − 1) (e λm M (x ) = ⎨ 2 tan φ (x − L) λm2 2 c P λ (m − m ) c P ⎪− ( − 1) − 2 tan φ 2 ( tan φ + λX ) ) ( tan φ + λX )(e λm2 2 tan φ ⎪ ⎪ (x − L) ⎪ ρ g (x − L)3 sin θ cos θ ⎪ + 2 6 ⎪ ⎪ (x ⩾ L) ⎩

N

Damage to hydraulic prop

(13) According to the relationship between the bending moment and the elastic deformation energy, the elastic deformation energy at any x along the rock pillar can be expressed as follows:

b N

3

Reversed loader Bending of Hydraulic prop

c

2

(ρ + ρ ) gx sin θ cos θ 1 ⎧ ⎡ 1 c ⎤ 2EI 6 ⎪ ⎣ ⎦ ⎪ (x ⩽ L) ⎪ 2 2 ⎪ ρ gL2 sin θ cos θ (x − L ) ρc gx 3 sin θ cos θ λm 2 c PX ⎤ 3 ⎪ ⎡ 1 + + + ( ) ( ) ⎥ ⎢ λ 2 6 2 tan φ tan φ ⎪ ⎥ ⎢ φ (x − L) ⎪ ⎪ ⎢ (e 2 tan λm ⎥ − 1) U (x ) = ⎥ ⎨ 1 ⎢ φ x L − 2 tan ( ) ⎥ λm2 2 c PX λ (m − m2) ⎪ 2EI ⎢ λm2 − + − − ( ) ( )( e 1) ⎥ λ 2 tan φ tan φ 2 tan φ ⎪ ⎢ ⎥ c PX ⎪ ⎢ + − ( )( x L ) ⎥ ⎢ ⎪ tan φ λ ⎥ ⎢ ⎪ 3 ρ g (x − L) sin θ cos θ ⎥ + 2 ⎪ ⎢ 6 ⎦ ⎪ ⎣ ⎪ (x ⩾ L) ⎩

N

(14) 1.3 m

3.3. Factors affecting the elastic deformation energy of the hard roof and hard-rock pillar Subsidence of upper of the roadway

Eqs. (8) and (14) show that the main factors influencing the elastic deformation energy of the roof and rock pillar are the elastic modulus, the moment of inertia, the angle of internal friction at the coal-rock interface, the cohesive force at the coal-rock interface, the dip angle of the coal seam, the supporting force coefficient, the lateral pressure coefficient, the thickness of the coal seam, the mining depth, the average density of the overlying strata and rock pillar, the density of the loess, and gravitational acceleration. The physico-mechanical parameters of the coal seam, loess, and rock strata have been determined according to the actual field conditions at the WCM, China (see Section 2.1), and are as follows: The elastic modulus, E, was 26.92 GPa. The angle of internal friction, φ, and the cohesive force, c, at the coal-rock interface were 30° and 1 MPa, respectively. The dip angle of the coal seam, θ, was 87°. The supporting force coefficient, k, was 0.5, and the lateral pressure coefficient, λ, was 0.5, while the thicknesses of the two coal seams that were simultaneously mined were m = 49.6 m and m2 = 36.5 m, and the mining depth, H, was 400 m. The average density of the rock strata, ρc, was 2,813 kg/m3. The density of the loess, ρ1, was 1,600 kg/m3, and the density of the coal, ρ2, was 1,336 kg/m3. Gravitational acceleration, g, was 9.8 N/kg. Since the following parameters are of high relevance for the scope of this study, the dip angle of the coal seam, the lateral pressure coefficient, and the supporting force coefficient have been selected as parameters for analyses so that their effects on the elastic deformation energy of the roof and rock pillar can be determined.

d N

Down wall

Up wall N

Side rollover of belt conveyor Direction of occurrences of the rock burst

Fig. 4. In situ photographs, recording the most disastrous rock bursts at working face No. B3 + 6. (a) Rock burst “3⋅13″, (b) rock burst ”11⋅24″, (c) rock burst “2⋅1″, and (d) rock burst ”4⋅26″ (rock burst designators refer to the month and day of their occurrence).

3.3.1. Dip angle of the coal seam As shown in Fig. 12(a), the elastic deformation energy of a roof of a given length increases with increasing dip angle of the coal seam 8

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Tailentry

Rock burst “4·26”

Headentry Tailentry

Rock burst “2·1”

Headentry Rock burst “11·24”

Tailentry Headentry Tailentry

Rock burst “3·13”

Headentry Tailentry

Rock burst “2·27”

Headentry

Fig. 5. Statistical analysis of the damage range of recorded rock bursts.

with increasing lateral pressure coefficient (see Fig. 13(a)). When λ = 0.5, the elastic deformation energy of a roof is in the range 400–450 m, which is 2 to 3 times higher than at λ = 1.1. This is because when the lateral pressure coefficient is small, the horizontal stress is relatively high, which tends to cause bending deformation of the roof of a coal seam with large dip angle. As a result, the elastic deformation energy of the overhanging hard roof increases, and the roof is easier to break when subjected to external disturbance, which is consistent with the results of Driad-Lebeau et al. (2005). The elastic deformation energy for a rock pillar of a given length increases with increasing lateral pressure coefficient (see Fig. 13(b)). This is mainly because a larger lateral pressure coefficient means that a smaller horizontal force acts on the left and right sides of the rock pillar. Hence, a smaller supporting force is given to the rock pillar, which results in its flexural deformation under the action of gravity and an increasing accumulation of elastic deformation energy.

between 0° and 72.6°, and then decreases as the dip angle continues to increase. With all other conditions being the same and for the same roof length, the roof in a steeply inclined coal seam is easier to break and more prone to rock burst under the action of intensive mining compared with a gently inclined or horizontal coal seam. This phenomenon explains why the depth at which rock burst occurs in the mining of inclined and steeply inclined coal seams is shallower than that in the mining of horizontal and gently inclined coal seams (Lai et al., 2015), and the critical mining depth for rock burst occurrence in coal seam No. B3 + 6 was only 300 m. Turning to the rock pillar, the elastic deformation energy in a rock pillar of a given length increases with increasing dip angle of the coal seam between 0° to 45°, and then decreases as the dip angle continues to increase (Fig. 12(b)). Furthermore, the elastic deformation energy of a roof of a certain length is much larger than that of a rock pillar of equal length. This is mainly because one side of the roof is directly subjected to horizontal tectonic stress and overlying load, while the other side contacts the goaf. Therefore, bending deformation of the roof easily occurs in the goaf direction. The rock pillar has goafs filled with loess on both sides. The horizontal stress and other forces must pass through the loess to reach the rock, thereby being considerably attenuated. The force acting on the rock pillar in the horizontal direction is small, and there is little difference between the left and right sides of the rock pillar. It is mainly bent by its own weight.

3.3.3. Supporting force coefficient The elastic deformation energy at a given roof or rock pillar length decreases with increasing supporting force coefficient (see Fig. 14). This is because a larger supporting force coefficient means that the loess filling the goaf is denser. This exerts a better supporting effect on the roof and rock pillar, the displacement and deformation of the roof and rock pillar along the goaf will be lower, and the stress concentration will also be lower. As a result, the elastic deformation energy of the roof and rock pillar decreases. Intensive mining leads to the instantaneous

Damage ranges of rock bursts (m)

3.3.2. Lateral pressure coefficient The elastic deformation energy in a roof of a given length decreases 500

>100

50-100

450

20~50

”20

400 350

42.1%

300 250 200

34.2%

150 100

18.4%

50

5.3%

0 0

20 40 60 Recorded rock burst events (times)

80

Proportions of rock bursts in different damage ranges

Fig. 6. Damage ranges of different rock burst events and their relative proportions in both horizontal and gently inclined coal seams in eastern China. 9

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Y

0

Loess

G

Fd

Coal seam

ș Fig. 9. Simplified mechanical model of the “steeply inclined suspended roof structure”.

X Fig. 7. Physical model of the “steeply inclined suspended roof structure”.

Fig. 10. Mechanical model of the “steeply inclined suspended rock pillar structure”.

higher than that of the above calculation results. Therefore, where the other parameters are difficult to change, the suspended roof and rock pillar should be artificially broken when the suspended roof and rock pillar reach a specific length so as to reduce the risk of rock burst.

Fig. 8. Mechanical model of the “steeply inclined suspended roof structure”.

4. Analysis of the dynamic and static stress

settlement of the loess filling the goaf and reduces the supporting force coefficient, and thus, the supporting force. The elastic deformation energy of the corresponding position on the roof and rock pillar increases, and rock burst is easily induced. Simultaneously, these results show that if the other factors remain fixed, the longer the length of the suspended roof and rock pillar, the greater the elastic deformation energy of the roof and rock pillar. Their elastic deformation energy reaches a maximum in the plastic zone, which is consistent with the location of the source of the rock burst described in Section 2.2. Simultaneously, Miao et al. (2016) reported that rock burst is likely to occur when the strain energy accumulated in the surrounding rock exceeds 1.0 × 105 J/m3. The strain energy accumulated in the roof and rock pillar during loss of support will be

4.1. Static stress induced by compressive and prying effects On the basis of the mechanical analysis in Section 3, the static stress of coal seam No. B3 + 6 has been simulated by FLAC3D to identify the degree of stress concentration. The model has a dimension of 600 m × 2500 m × 500 m. The dimensions of the coal seam and rock pillar are the same as in the actual in-situ situation. The strain-softening model has been adopted in the filling of the goaf, and the Mohr-Coulomb failure criterion has been adopted in the coal and rock mass. The mechanical parameters used in this model have been obtained via laboratory experiments. The horizontal stress of the working face position when the rock burst “4⋅26” occurred is shown in Fig. 15. It can be seen 10

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Fig. 11. Simplified mechanical model of the “steeply inclined suspended rock pillar structure”.

Fig. 13. Variation law of elastic deformation energy with different lateral pressure coefficients: (a) roof and (b) rock pillar.

energy, as shown in Section 3. This indicates that when the suspended roof and rock pillar bend to the goaf, the coal body will be subjected to the coupled action of the compressive and prying effects of the roof and rock pillar. This will increase the horizontal stress of the coal body at the compressive and prying area beyond the horizontal tectonic stress, and will lead to stress concentration in the compressive and prying areas. Simultaneously, it should be emphasized that the above simulation results have been obtained based on the condition of full goaf filling. If a fully mechanized caving mining method is used to extract coal at a height of 25 m, and the subsidence of the goaf infill is induced, the compressive and prying effects will be more obvious. 4.2. Dynamic stress induced by hard roof and hard-rock pillar breakage MS events induced by roof and rock pillar breakage will produce dynamic stress in surrounding coal and rock mass. The distribution of MS events is closely related to the degree of stress concentration (Orlecka-Sikora, 2010). As described in Section 2.3, the proportions of MS sources on the hard roof and hard-rock pillar are 17.0% and 60%, respectively. The intensive occurrence of MS events in the roof indicates that the goaf above the mining level placed the roof above the mining level in a suspended state. The combined action of the overlying strata and high horizontal tectonic stress, produces bending, deformation, and stress concentration in that part of the roof. It also causes the roof to accumulate a large amount of elastic deformation energy, thus resulting in the occurrence of MS events under external disturbance. The rock pillar is in a suspended state as the coal seams on both sides continue to

Fig. 12. (a) Distribution of the elastic deformation energy of the roof for coal seams with different dip angles, and (b) distribution of elastic deformation energy of the rock pillar for coal seams with different dip angles.

that the peak stress of coal seam No. B3 + 6 was 29.6 MPa, located 25 m below the roadway, which was 1.7 times the horizontal tectonic stress at this level, 2.4 times the uniaxial compressive strength, and 24.4 times the tensile strength of the coal mass. The vertical position of the peak stress in the roof and rock pillar is similar to the position of the peak stress in the coal seam. This position is highly consistent with the position of the roof and rock pillar with high elastic deformation 11

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be mined to deeper depths. Under the combined action of horizontal stress under the mining level, the weight of the rock pillar and mining disturbance, the rock pillar becomes unstable and suffers a rotary prying effect. Stress concentration occurs at the rotating and prying positions (Fig. 15), which accumulate a large amount of elastic deformation energy, resulting in the intensive occurrence of MS events. With mining disturbance and instantaneous fall of goaf fill the supporting force coefficient is reduced as is the supporting force. The strain energy accumulated in the roof and rock pillar during loss of support will be higher than that of the energy calculation in Section 3. As such, breakage occurs easily under external disturbance to make more MS events. As shown in Fig. 3(a), the energy of rock bursts recorded since the application of the MS monitoring system are 5 × 108 J, 9.5 × 106 J, 2.1 × 108 J, and 2.2 × 106 J, respectively. The distances between the source and the coal-rock interface were 34.9 m, 12 m, 9 m, and 35 m, respectively. MS monitoring shows that the velocity of particle vibration increases with increasing MS energy (He et al., 2015). The particle vibration velocity reaches a maximum at the source boundary and attenuates gradually as it moves away from the source. The distance between the MS source boundary and the source center can be estimated by the following equations (Hatherly et al., 2003)

r1 =

3

2W1 G π Δτ 2

(15)

where r1 represents the shear fracture radius of coal-rock, r2 represents the tensile fracture radius of coal-rock; Δτ and Δσ represent the stress drop caused by shear and tensile fracture, respectively, which, according to the Mohr-Coulomb and Griffith strength criteria, are generally 1/2 and 1/8 of the uniaxial compressive strength, respectively; W1 represents the strain energy released by shear fracture, W2 represents the strain energy released by tensile fracture; G represents the average shear modulus of coal-rock, and E represents the average E elastic modulus of coal-rock. Furthermore, G = 2(1 + μ) , where μ represents Poisson's ratio. Laboratory tests have shown that the average modulus of elasticity and Poisson's ratio of the roof and rock pillar were 23.43 GPa and 0.25, respectively. The uniaxial compressive strength was 51.82 MPa, and the density was 2500 kg/m3. For the 2.2 × 106 J rock burst mentioned above, the rupture radius corresponding to shear rupture of 2.7 m, and that corresponding to tensile rupture was 7.29 m. Li (2016) found that

Fig. 14. Variation law of elastic deformation energy with different supporting force coefficients: (a) roof and (b) rock pillar.

Roof

No. B3+6

Rock pillar

No. B1+2

+475 Working face

+450

Prying region

Compressive region

Fig. 15. Stress distribution in coal and rock mass. 12

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when the MS event energy is at the 106 J level, the particle vibration velocity at the source boundary is 8.45–12.27 m/s, the P-wave velocity and S-wave velocity of the complete rock mass are 4000 m/s and 2300 m/s, respectively, and the attenuation coefficient is 1.526. The energy of the rock burst “4⋅26″ is 2.2 × 106 J, and this study assumed that v0 was 8.45 m/s. The maximum normal and tangential dynamic stresses generated by the rock burst ”4⋅26″ can be calculated by the following equations (modified according to Li, 2016):

of meters to hundreds of meters. Due to the high strength, large thickness, and good integrity of the hard roof, a considerable amount of elastic energy can be stored in the roof under the combined action of the horizontal tectonic stress and the overlying load. Simultaneously, the horizontal tectonic stress and the overlying load have been applied to the coal body through the suspended roof, which results in a stress concentration in the corresponding position in the coal body. Strong mining disturbance and instantaneous subsidence of the filling in the goaf can cause a certain length of the suspended roof to break because the strain energy accumulated in the roof during loss of support is higher than that of the energy calculated in Section 3. The stored elastic energy can be released abruptly, which may induce dynamic disasters, such as rock burst (Huang et al., 2017; Huang et al., 2018). The suspended roof breakage (dynamic stress) and high static stress in the coal seam are the main factors that lead to rock burst, as shown in Fig. 18. This mechanism can be used to explain the cause and damage characteristics of the rock burst “2⋅1″ as described in Section 2.3.

j

⎧ σ = ρ C v ∏ l −αi i j Pj P 0 ⎪ dP i=1 j ⎨ −αi ⎪ σdS = ρj CSj vS 0 ∏ li i=1 ⎩

(16)

where li represents the propagation distance of the vibration wave in the i-th medium; αi represents the attenuation coefficient of the i-th medium; σdP and σdS represent the normal and tangential stresses caused by the propagation of the vibration wave to lj m away from the source boundary, respectively; ρj represents the density of the medium lj m away from the source boundary; CPj and CSj represent the propagation velocity of P waves and S waves in the medium lj m away from the

5.2. Rock pillar prying and breakage-induced rock bursts

j

As the coal seams on both sides of the rock pillar continue to be mined to deeper depths, the top coal mining causes the filling in the goaf to sink instantaneously, thus leading the rock pillar to become unstable and bending in the direction of inclination. This results in a stress concentration in the area of the coal-body prying and the corresponding area of the rock pillar. The larger the degree of bending, the more elastic deformation energy is accumulated in the rock pillar. The strain energy accumulated in the rock pillar during bending is higher than that of the energy calculated in Section 3. When the limit of the energy storage is reached, the rock pillar breaks, triggered by external disturbance, and the stored energy is released instantaneously, which may induce a rock burst. The instability and breakage of the rock pillar (dynamic stress) and the high static stress in the prying area of the coal seam are the main factors that lead to the rock burst, as shown in Fig. 19. This mechanism can be used to explain the cause and damage characteristics of other rock bursts as described in Section 2.3.

source boundary, respectively; l j = ∑ li , and j represent the j-th proi=1

pagation medium; vP0 and vS0 represent the particle vibration velocity of P waves and S waves at the source boundary, respectively. According to Eq. (16), the maximum normal and tangential dynamic stresses generated by the rock burst “4⋅26″ are 84.5 MPa and 48.6 MPa, respectively. The energies of the other rock bursts are greater, and the resultant dynamic stresses are higher. Thus, the dynamic stress caused by a high-energy tremor induced by roof and rock pillar breakage has considerable destructive force. Based on the analysis result of elastic deformation energy in Section 3 and static stress in Section 4.1, if the energy released from the roof and rock pillar breakage reaches a certain level and propagates to the position where the coal body is subjected to compression and prying action, the possibility of induced rock burst is higher. 5. Rock burst mechanisms Based on the MS monitoring data, the characteristics of the rock burst damage, the results of the mechanical model analysis, the numerical simulation, and the dynamic analysis. The rock burst-causing process in coal seam No. B3 + 6 was determined is shown in Figs. 16 and 17. The damage model for the roof and rock pillar, and the damage process are shown in Fig. 16. The damage model for the coal pillar, and the damage process are shown in Fig. 17. Fig. 16 shows that the indirect causes of the induced rock burst are (1) the high horizontal tectonic stress, (2) subsidence of the goaf infill and lower supporting force coefficient, (3) instability of the roof and rock pillar and accumulation of elastic deformation energy, (4) elastic deformation energy of roof and rock pillar at corresponding position exceeds the energy storage limit, and (5) breakage of the roof and rock pillar. Fig. 17 shows that the immediate cause of the induced rock burst is that the critical stress required for rock burst occurrence is exceeded by the superposed dynamic and static stresses. The static stress originates from the tectonic stress applied to the coal body and the compressive and prying effects of the roof and rock pillar. The wave generated by roof/rock pillar breakage transfers to the coal pillar and acts on the coal pillar in the form of dynamic stress. Therefore, three mechanisms by which rock bursts can occur in SIETCS are proposed (Sections 5.1–5.3).

5.3. Rock burst induced by coupling of compressive and prying action In addition to the above cases (Sections 5.1–5.2), there must be a further case. In this case, the superposition of static stress in the coal seam caused by the compression from the suspended roof and dynamic stress caused by the breakage of the roof has not quite reached the critical level. Moreover, the superposition of the static stress in the coal seam caused by prying of the rock pillar and the dynamic stress caused by the breakage of the rock pillar alone are not sufficient to induce a rock burst. However, the simultaneous action of both factors can lead to a rock burst. The dynamic stress caused by the breakage of the roof and rock pillar and the static stress of the coal seam under the combined action of the compressive and prying effects can superpose so that the total stress meets the stress criterion for a rock burst, thus promoting its occurrence. The breakage of the roof and rock pillar and the compressive and prying effects are the main burst-inducing factors for the rock burst as shown in Fig. 20. Since coal seam No. B3 + 6 is clamped by its roof and a rock pillar, it forms a “Roof–B3 + 6 coal seam–Rock pillar” structure, and the coal seam is simultaneously subjected to the large load and prying action. Breakage in one component triggers changes in the others and induces a rock burst under the coupled effect of multiple factors. Judging from the damage characteristics of rock bursts occurred in coal seam No. B3 + 6, most rock bursts are the result of the combined action of the roof and rock pillar rather than being caused by a single factor only.

5.1. Suspended roof compression and breakage-induced rock bursts The length of the suspended roof increases continuously with increasing mining depth. The suspended roof then appears in the goaf above the mining level, and the suspended length becomes several tens 13

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a

Bending of roof and rock pillar

Subsidence of the goaf infill

High horizontal tectonic stress

Rock pillar Goaf

Goaf

+500

B1+2 Breakage

+450

Prying region

B3+6

Compressive region

+475

Wave

b Fully mechanized caving mining

Instantaneous subsidence of the goaf infill

Roof/rock pillar loss of support

Breakage of the hard roof/rock pillar

Accumulation of elastic deformation energy and exceeds the energy storage limit

Instability of the roof and rock pillar, and this produces the compressive and prying effects on the coal body

Fig. 16. Schematic diagram of the damage of both roof and rock pillar. (a) Damage model for roof and rock pillar, and (b) damage process in roof and rock pillar.

6. Principles and effects of rock burst prevention engineering

propagates to this area of stress concentration and superposes the static stress, the strength limit of the coal body is exceeded, and a rock burst is induced. The rock pillar is suspended between both coal seams with a dip angle of 87°. Rotary instability is prone to occur under external disturbance, thus producing a prying effect on coal seam No. B3 + 6 and resulting in a high static stress concentration in the prying area. Subject to strong mining disturbance, the rock pillar is easily destabilized and fails, thus resulting in a dynamic load disturbance, which superposes with the high static stress and induces a rock burst. Section 3 shows that the suspended length of the roof and the rock pillar, the lateral pressure coefficient, and the supporting force coefficient exert important influences on the elastic deformation energy of the roof and rock pillar. The risk of rock burst in SIETCS can be reduced by reducing the suspended length of the roof and rock pillar or the horizontal tectonic stress applied to the coal seam, or by increasing the density of goaf infill. However, the density of goaf infill is difficult to change. Therefore, a load-reduction and prying-reduction project plays an important role in the prevention of rock burst in SIETCS.

6.1. Load-reduction and prying-reduction project On the basis of the above research results, it is concluded that the tectonic stress and a hard, suspended roof and rock pillar provide the main sources of the forces that lead to the occurrence of a rock burst in the SIETCS at the WCM. Tectonic stress is a major cause of coal-rock mass failure and rock bursts (Yan et al., 2016; Wang et al., 2016). For example, the rock burst on October 15, 2012, in the Xing’an Coal mine, Hegang City, China, occurred in a synclinal region and caused severe damage to the 104 m gateway. Coalface 3431 at Huating coal mine, located in a large fold, suffered ten rock bursts from March 1, 2014, to April 8, 2014, during which, the coalface advanced a mere 80 m. At the WCM, the high horizontal tectonic stress and the overlying strata act on the coal seam through the suspended roof in the form of a load. This results in a stress concentration in the corresponding area in the coal body. When dynamic load stress, caused by the breakage of the roof, 14

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a

b

Fig. 17. Schematic diagram of the damage in the coal pillar. (a) Damage model for the coal pillar at compressive and prying region, and (b) damage process in the coal pillar. σs represents the static stress, σd represents the dynamic stress, and σc represents the critical stress when rockburst occurs.

pressure relief is conducted, it is restricted by the sealing conditions, and the rock mass structure in the deep area is destroyed while the shallow area is not, thus tending to form the “plate” structure shown in Fig. 21. Under the influence of an increase in mining depth and mining disturbance, tension cracks appear in the roof due to the combined action of weight and tectonic stress, and the rock mass on the upper part of the “plate” structure slips in the goaf direction and acts on the “plate” structure. The “plate” structure thus becomes the carrier of energy accumulation. Mining disturbance causes the “plate” structure to break instantaneously, and the ultimate energy accumulated in the “plate” structure is suddenly released, which induces a rock burst. Therefore, in this study, alternate deep and shallow hole blasting modes have been selected to break up the roof and to achieve load-reduction. The technical parameters of the blasting holes used for the roof construction were as follows: The row spacing of shallow blasting holes was 10 m with three holes in each row with construction angles of 25°, 45°, and 65°. The borehole length was 25 m or 30 m, and the borehole

Furthermore, based on the research results in Section 2.3, a project at the WCM should be carried out within 209 m ahead of the working face. 6.1.1. Load-reduction project The analysis of elastic deformation energy indicates that destroying the integrity of the roof and forming a three-dimensional buffer zone at the junction of the roof and the coal body are effective engineering measures. These reduce the transfer of horizontal stress and overlying load from the roof and lateral pressure coefficient and the accumulation of elastic deformation energy in the roof and compressive stress in the coal body. Destress blasting techniques have been successfully employed in underground mines with the aim of preconditioning a highly stressed rock mass to mitigate the risk of rock burst occurrence in mines (Sainoki et al., 2017; Vennes and Mitri, 2017). A large number of blasting engineering studies at the WCM research site showed that shallow hole blasting does not achieve a sizeable pressure relief effect on the deep rock mass. When deep hole blasting 15

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diameter was 113 mm. The row spacing of the deep blasting holes was 10 m, with two holes in each row with construction angles of 25° and 45°. The borehole length was 50 m, and the borehole diameter was 113 mm. Loess was used to seal the holes. The setting of different drilling angles and hole depths aims to increase the pressure relief range and ensure effective pressure relief. A diagram of the load-reduction project is shown in Fig. 22. 6.1.2. Prying-reduction project The prying effect can be reduced by destroying the integrity of the interlayer rock pillar and by reducing the overall height of the rock pillar over time. Simultaneously, these measures can also reduce the transmission of horizontal stress to the mining face; this plays an important role in reducing the rock burst risk at SIETCS. The integrity of the rock pillar is destroyed by combining alternately blasting deep and shallow holes, water injection, and drilling hole blasting. The technical parameters of the prying-reduction project for the rock pillar were as follows: The parameters of the deep and shallow holes are consistent with the technical parameters of the load-reduction project. The length, width, and height of the drilling field were 5 m, 6 m, and 2.5 m, respectively. The distance between the two drilling fields was 300 m, the length of the water injection hole was 135 m, and the diameter of the borehole was 113 mm. The specific technical parameters and the layout of the prying-reduction project are shown in Fig. 23.

Fig. 18. Sketch showing the mechanism associated with suspended roof compression and breakage-induced rock burst.

6.2. Effectiveness of preventative engineering 6.2.1. Comparison of AE monitoring results before and after the implementation of the load-reduction and prying-reduction project The KJ623 AE monitoring system was installed in the WCM to calculate the occurrence time, energy released, and number of pulses. The system contains a total of 11 sensors that were installed on the coal and rock mass to collect AE events within a frequency range of 300–2000 Hz. A bolt has been used as AE conductor, and the sensor was installed on the bolt. The sensors have been arranged 30 m away from the working face with a spacing between each sensor of 30–40 m; each sensor covers a range of 50 m. To facilitate the comparison of AE energy and the number of pulses before and after the implementation of the load-reduction and prying-reduction project, the parameters “millionton AE energy” AE per and “million-ton AE pulse number” AP per have been defined as the AE energy and number of pulses recorded for each one million tons of coal produced at the working face. The expressions for AE per and AP per are as follows:

Fig. 19. Sketch showing the mechanism associated with rock pillar prying and breakage-induced rock bursts.

AE per =

ET × 10−6 LW LL hρ2

(17)

AP per =

FT × 10−6 LW LL hρ2

(18)

where ET represents the total energy of AE events during working face mining, PT represents the total number of AE pulses during working face mining, LW represents the width of the fully mechanized top-coal caving face, LL represents the length advanced at the fully mechanized top-coal caving face, h represents the height of the fully mechanized top-coal caving face, and ρ represents the density of the coal, ρ2 = 1336 kg/m3. The AE monitoring results before the implementation of load-reduction and prying-reduction (i.e. from December 1, 2016, to March 13, 2017) and after the implementation of load-reduction and pryingreduction (i.e. from June 12, 2017, to March 22, 2018) have been analyzed. Before and after the implementation of load-reduction and prying-reduction, AE per were 14.5 × 106 J and 8.7 × 106 J, respectively, and AP per were 10.7 × 106 and 11.2 × 106, respectively. A comparison of before and after AE per shows that AE per decreased by 40%, and a comparison of before and after AP per shows that AP per has increased by 4.5%. The increase in AP per is closely related to the effects

Fig. 20. Sketch showing the rock burst-inducing mechanism involving coupled compressive and prying effects.

16

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Fig. 21. (a) Schematic diagram showing the horizontal tectonic stress and gravitational force from overlying strata acting on a “plate” structure, and (b) in situ photographs, recording the roof split.

implementation of the load-reduction and prying-reduction project. The longer project implementation, the more the pressure was relieved, and the more the daily total energy decreased. The spatial distribution of MS sources before and after project implementation is shown in Fig. 25. After the implementation of load-reduction and prying-reduction, there was a marked reduction in the number of MS events in the roof and interlayer rock pillar; and the longer the project was implemented, the fewer MS events were recorded. Before the implementation of the project, there were, per-meter, 0.04 MS events with an energy above 105 J, 0.69 MS events with an energy between 104–105 J, and 3.16 MS events with an energy between 103–104 J. After the implementation of the project, the number of MS events per-meter were 0.01 with an energy above 105 J, 0.02 MS events with an energy between 104–105 J,

of the pressure relief project, which destroys the integrity of the rock mass and increases the number of micro-cracks. The results show that this measure can reduce the stress concentration of the coal and rock masses in SIETCS, decrease the dynamic load disturbance, and reduce the risk of rock burst.

6.2.2. Comparison of MS monitoring results before and after the implementation of the load-reduction and prying-reduction project The changes in the daily total MS energy before and after the implementation of the load-reduction and prying-reduction project are shown in Fig. 24. Stage “A” is before the implementation of the project, and stage “B” is after the implementation of the project. As shown in Fig. 24, the daily total energy decreases as a whole after the 17

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Fig. 22. Schematic diagram of the load-reduction project: (a) schematic diagram of the load-reduction principle, (b) layout of the load-reduction project, (c) section plan of shallow blasting holes, and (d) section plan of deep blasting holes.

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Fig. 23. Schematic diagram of the prying-reduction project: (a) layout of the prying-reduction project, and (b) elevation view of the layout of blasting holes in the drilling field.

seam and bending deformation of the rock pillar are also reduced. The above research results show that these load-reduction and prying-reduction measures are effective in reducing the risk of rock burst in a SIETCS. 6.2.3. Comparison of roadway deformation before and after the implementation of the load-reduction and prying-reduction project A total of three rock bursts occurred in coal seam No. B3 + 6 at the + 450 horizontal working face before the implementation of the load-reduction and prying-reduction project, and the roadway was severely damaged. After the implementation of preventative measures, no further rock bursts occurred in coal seam No. B3 + 6 during mining. The phenomena of wall movement and floor heave in the roadway were also clearly reduced. This reduction of damage to the roadway can be attributed to the weakening of the dynamic and static loads by the loadreduction and prying-reduction project.

Fig. 24. Curves showing the variation in the daily total MS energy before and after the implementation of the load-reduction and prying-reduction project. A: Before the implementation of the project; B: after the implementation of the project.

7. Conclusions

and 1.18 MS events with an energy between 103–104 J. The main reasons for the above results are that the load-reduction and prying reduction measures destroy the integrity of the roof and rock pillar, reduce the length of the suspension, change the values of the lateral pressure coefficient and supporting force coefficient, weaken the ability of the roof to transfer horizontal tectonic stress and weight from the overlying strata, and decrease the elastic deformation energy accumulated in the roof and rock pillar. Furthermore, prying on the coal

This study analyzed the mechanism and preventative principle of rock burst under the condition that two adjacent steeply inclined coal seams are mined simultaneously to a deep depth under the condition of goaf filling. The central findings are listed below: (1) The rock burst damage at the WCM usually show clear directionality. The damage range of the rock burst is mostly concentrated 19

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Fig. 25. (a) Spatial distribution of MS events with energy exceeding 1.0 × 103 J before the implementation of the load-reduction and prying-reduction project, (b) spatial distribution of MS events with energy exceeding 1.0 × 103 J after the implementation of the load-reduction and prying-reduction project.

recorded rock bursts were 84.5 MPa and 48.6 MPa, respectively. The main factors causing rock bursts in SIETCS are dynamic stressinduced breakage of the suspended roof and rock pillar and static stress under the combined action of compressive and prying forces. The forces that induced rock bursts are mainly imposed by the roof, interlayer rock pillar, and horizontal tectonic stress. (4) The damage models and damage process by which rock burst is induced have been developed. Three mechanisms by which rock burst can occur in SIETCS are presented in this paper, and the prevention principles of load-reduction and prying-reduction for rock bursts in SIETCS have been introduced. The application of a load-reduction and prying-reduction project has been successfully demonstrated in the WCM.

0–209 m in front of the working face. The rupture zones of the rock mass are mainly concentrated in the roof and interlayer rock pillar, and the proportions of MS sources that have been recorded in the roof and rock pillar are 17%, and 60%, respectively. (2) The mechanical models of a “steeply inclined suspended roof structure” and a “steeply inclined suspended rock pillar structure” have been developed, and the elastic deformation energy distribution functions for both models have been obtained. The elastic deformation energy of the roof and rock pillar are closely related to the dip angle of the coal seam, the lateral pressure coefficient, the supporting force coefficient, and the length of the suspended roof and rock pillar. The risk of a rock burst in a SIETCS can be reduced by controlling these parameters. (3) The peak stress of the coal body at the compressive and prying area was 1.7 times the horizontal tectonic stress. The minimum normal and tangential dynamic load stresses generated by previously 20

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Conflicts of interest

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The authors confirm that there are no conflicts of interest associated with this publication. CRediT authorship contribution statement Shengquan He: Conceptualization, Methodology, Data curation, Writing original draft, Visualization. Dazhao Song: Conceptualization, Writing - original draft, Writing - review & editing, Supervision, Funding acquisition. Xueqiu He: Validation, Supervision, Funding acquisition. Jianqiang Chen: Investigation, Data curation. Ting Ren: Writing - review & editing, Visualization. Zhenlei Li: Methodology, Investigation. Liming Qiu: Validation, Visualization. Acknowledgments Special thanks is extended to the Wudong coal mine for the provision of raw data. This work was financially supported by the State Key Research Development Programme of China (No. 2016YFC0801408), the National Natural Science Foundation of China (Nos. 51634001 and 51774023), the Fundamental Research Funds for the Central Universities (FRF-TP-18-007C1), and the China Postdoctoral Science Foundation (2018M641201). References Blake, W., Hedley, D.G.F., 2009. Rockbursts: case studies from North American hard-rock mines. Society for Mining Metal-lurgy and Exploration Inc, Littleton CO. Cai, W., Dou, L.M., Zhang, M., Cao, W.Z., Shi, J.Q., Feng, L.F., 2018. A fuzzy comprehensive evaluation methodology for rock burst forecasting using microseismic monitoring. Tunn. Undergr. Space Technol. 80, 232–245. Calleja, J., Porter, I., 2016. Coalburst control methods. Coal Operators' Conference. University of Wollongong, Australia. Chen, G.X., Dou, L.M., Xu, X., 2012. Research on prevention of rock burst with relieving shot in roof. Proced. Eng. 45, 904–909. Chen, X.J., Li, L.Y., Wang, L., Qi, L.L., 2019. The current situation and prevention and control countermeasures for typical dynamic disasters in kilometer-deep mines in China. Saf. Sci. 115, 229–336. Cook, N.G.W., 1963. The seismic location of rockbursts. In: Proceedings 5th Symp. Rock Mech. Pergamon Press, pp. 493–516. Deng, Y.H., Wang, S.Q., 2014. Feasibility analysis of gob-side entry retaining on a working face in a steep coal seam. Int. J. Min. Sci. Technol. 24 (4), 499–503. Dou, L.M., He, X.Q., 2001. Prevention theory and technology of rockburst. China University of Mining and Technology Press, Xuzhou. Dou, L.M., Mu, Z.L., Li, Z.L., Cao, A.Y., Gong, S.Y., 2014. Research progress of monitoring, forecasting, and prevention of rockburst in underground coal mining in China. Int. J. Coal Sci. Technol. 1, 278–288. Díez, R.R., Álvarez, J.T., 2000. Hypothesis of the multiple subsidence trough related to very steep and vertical coal seams and its prediction through profile functions. Geotech. Geol. Eng. 18 (4), 289–311. Driad-Lebeau, L., Lahaie, F., Heib, M.A., Josien, J.P., Bigarré, P., Noirel, J.F., 2005. Seismic and geotechnical investigations following a rockburst in a complex French mining district. Int. J. Coal Geol. 64 (1–2), 66–78. Feng, G.L., Feng, X.T., Xiao, Y.X., Yao, Z.B., Hu, L., Niu, W.J., Li, T., 2019. Characteristic microseismicity during the development process of intermittent rockburst in a deep railway tunnel. Int. J. Rock Mech. Min. Sci. 124, 1–13. Feng, X.J., Wang, E.Y., Shen, R.X., Wei, M.Y., Chen, Y., Cao, X.Q., 2011. The dynamic impact of rock burst induced by the fracture of the thick and hard key stratum. Proced. Eng. 26, 457–465. Hatherly, P., Galel, M., Medhurst, T., 2003. “3D Stress effects, rock damage and longwall caving as revealed by microseismic monitoring”, ACARP Project C9021. ACARP, Brisbane, Australia. He, J., Dou, L.M., Cai, W., Li, Z.L., Ding, Y.L., 2015. In situ test study of characteristics of coal mining dynamic load. Shock Vib. 2015 (1), 1–8. He, J., Dou, L.M., Cao, A.Y., Gong, S.Y., Lv, J.W., 2012. Rock burst induced by roof breakage and its prevention. J. Cent. South. Univ. 19, 1086–1091. He, J., Dou, L.M., Gong, S.Y., Li, J., Ma, Z.Q., 2017. Rock burst assessment and prediction by dynamic and static stress analysis based on micro-seismic monitoring. Int. J. Rock Mech. Min. Sci. 93, 46–53. He, S.Q., Song, D.Z., Li, Z.L., He, X.Q., Chen, J.Q., Li, D.H., Tian, X.H., 2019. Precursor of spatio-temporal evolution law of MS and AE activities for rock burst warning in steeply-inclined and extremely-thick coal seams under caving mining conditions. Rock Mech. Rock Eng. 52, 1–21. Heib, M.A., 2012. Numerical and geophysical tools applied for the prediction of mine induced seismicity in french coalmines. Int. J. Geosci. 3 (24), 834–846. Hou, C.J., Ma, N.J., 1989. Stress in in-seam roadway sides and limit equilibrium zone. J.

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