# A Control Method of Rock Burst for Dynamic Roadway Floor in Deep Mining Mine

Authors: Zhimin Xiao; Jun Liu; Shitan Gu; Mingqing Liu; Futian Zhao; Yue Wang; Chen Ou; Mengyang Zhen
Journal: Shock and Vibration
 (2019)

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## Abstract

Roadway floor rock burst is an important manifestation of rock bursts in deeply buried mines. With the increase of mining depth and mining intensity, rock burst disasters in the roadway floor such as floor heaves are becoming more serious. The article investigated the roadway floor severe heave caused by floor rock burst during excavation of the No. 3401 working face, which was controlled by an anticlinal structure and deep mining in Shandong Mine, China. Firstly, by analyzing geological conditions of the working face, roadway support parameters, and characteristics of coal and rock, it was revealed that high tectonic stress and high crustal stress were main causes of the floor rock burst. Secondly, based on the Theory of Mechanics and Theory of Energy, the energy conversion process in the roadway floor was discussed, and the rock burst condition caused by elastic energy in the roadway floor was analyzed. The failure characteristics of roadway-surrounding rock were also inspected, using a borehole recorder. The roof and sidewalls of roadway mainly contained fissures and cracks, whereas cracks and broken areas are distributed in the roadway floor. Finally, based on the deformation and failure characteristics of roadway-surrounding rock, a method termed “overbreaking-bolting and grouting-backfill” was proposed to control roadway floor rock burst. The method was tested in the field, and the results showed that it could effectively control the deformation of roadway floor and rock burst, guaranteeing the stability of roadway floor. This impact control method for the roadway floor can provide a reference for the prevention and control of roadway rock burst in mines with similar geological conditions.


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## Body

## 1. Introduction

Rock burst, a special form of mine stress, can be described as an instantaneous release of high-strength coal (rock) strain elastic energy induced by mining. Vibrations in the surrounding rock and coal-rock extruding phenomenon occur in the corresponding mining space [1]. Since 1738, rock bursts have been reported in the leading coal mining countries, including the United Kingdom, Poland, South Africa, Russia, United States, and China [2–4]. With the increase of mining depths, frequency of rock bursts in mines increases significantly and the severity of resulting disasters is also increasing. According to the available statistics, there are more than 130 mines in China, and hundreds of casualties die every year from rock burst [5]. An investigation showed that 75% of rock burst accidents occurred in roadways [6, 7]. Consequently, the mechanism and control techniques of rock burst have been studied extensively by experts and scholars.Based on the mechanical characteristics of coal rock and characteristics of rock burst, many experts have studied the disaster from three angles: the mechanism, early warning, and prevention and control.Many scholars have conducted in-depth research on the mechanism and prevention of rock burst and made remarkable achievements. A series of theories, such as the Energy Theory, Stiffness Theory, Strength Theory, and Impact Tendencies Theory, have been put forward [8–11]. With the development of modern mathematics and computer technology, rock burst theories began to be investigated at the end of the 1970s. Based on the catastrophe theory [12–14], the damage propagation capacity of coal and rock mass was studied and the mechanism of rock burst was explored. Xie et al. [15] introduced fractal geometry into the study of rock burst, while Qi et al. [16] proposed the “three-factor” theory for studying the relationship between rock burst and coal-rock sliding failure. In the roadway rock burst theory, Ma et al. [17] discussed the mechanism and conditions of “butterfly” rock burst and put forward the “three criteria of rock burst,” providing a new method for its prevention and control. Yang et al. [18] obtained that passive acoustic emissions monitoring systems coupled with double-difference tomography were capable of monitoring and forecasting failure of rock, for helping improve the safety and efficiency of mining projects. Jiang et al. [19] obtained that the intermediate principal stress has a significant effect on rock burst behavior of marble. Xu et al. [20] discussed the characteristics of roadway rock burst and put forward the concept of “roadway floor risk factor.” Gu et al. [21] studied on rock burst risk induced by mining and field case.Rock burst warning is an important part of rock burst prevention and control, and many scholars have conducted long-term research and put forward a variety of rock burst danger prediction and forecasting methods. At present, early-warning methods for rock burst mainly include the microseismic method, electromagnetic radiation method, acoustic emission method, and borehole method [22–24]. On the topic of microseismic monitoring and early warning, Husen et al. [25] discussed the seismicity caused by stress redistribution during the excavation of the MFS Faido, the southernmost access point in the new Gotthard Base Tunnel. Zhao et al. [26] studied the characteristics of rock burst microseismic signals under a hard coal seam and hard roof. Sharp decreases in microseismic frequency and an obvious increase in amplitude were regarded as characteristic of rock burst precursors. Cao et al. [27] studied the microseismic precursory characteristics of rock burst hazard and assessed rock burst risk and rock burst risk areas in coal mining. On electromagnetic emission forecasting technology, He et al. [28] used electromagnetic forecasting’s pulse number and intensity as two indicators to classify danger of dynamical disasters into three levels of forecasting such as no danger, weak danger, and strong danger.With regard to the prevention and control of rock burst, Pan et al. [29] put forward two roadway support designs to improve support stiffness and energy absorption. Kang et al. [30] insisted that the essence of rock burst roadway support is maintaining the integrity of the surrounding rock, forming a supporting stress field and bringing into play the anti-impact capacity of the surrounding rock. Li et al. [31] proposed an optimized roadway layout for effective control of gob-side rock bursts in the longwall mining of a thick coal seam. Raju et al. [32] proposed methodology for designing dynamic rock supports by estimating the stored energy around the Canada Mine. Feng et al. [33] presented coal burst mitigation strategy for tailgate during deep mining of inclined longwall top coal caving panels by analyzing field data and numerical simulation results.However, with the increasing depth of coal mining, deep coal and rock mass are exposed by mining activities, and it can show complex geological characteristics and mechanical properties. Therefore, mining activities are more likely to induce severe and abrupt rock burst. In the article, the No. 3401 working face (in Shandong Mine, China) roadway floor heave induced by rock burst was studied as an example. First, the stability and energy laws of roadway floor were studied by using theoretical analysis and microseismic monitoring. Then, the internal failure characteristics of roadway-surrounding rock are determined and studied by using a borehole recorder. Finally, according to the deformation and failure characteristics of roadway floor, the “overbreaking-bolting and grouting-backfill” method was proposed. This method can improve the lithology of roadway floor, thus effectively controlling roadway floor deformation and guaranteeing roadway stability.


## 2. Rock Burst Manifestation of the Roadway Floor

### 2.1. Engineering Geological Characteristics of Working Face

The No. 3401 working face is the first working face in the No. 4 mining area in Shandong Mine, China, and the mining seam belongs to No. 3 coal seam. The working face elevation is −986 m to −914 m, corresponding to the ground elevation of +52.86 m to +54.75 m, and the average buried depth of working face is 1000 m, which is classified as a deep mining face. The average thickness of coal seam is 4.7 m, and it is affected by an anticlinal structure. The working face layout is shown in Figure1. Based on the 15 boreholes in the mine fine-searching stage, macroscopic structure of the No. 3 coal seam is plotted in Figure 2. These boreholes only have boreholes No. T9-6 and No. T10-3 passing through No. 3401 working face.Figure 1
Layout of No. 3401 working face.Figure 2
Macroscopic storage structure profile of No. 3 coal seam.Figures1 and 2 show that overall occurrence status of the No. 3 coal mine is controlled by the fold structure. The anticlinal structure greatly influences the No. 3 coal seam uplift, while the influence of synclinal structure is relatively small. The coal seam in the No. 3401 working face is greatly influenced by the anticline, and the geostress is large. Therefore, roadway excavation activity further increases the possibility of rock burst. Table 1 shows the results of a geostress test in the mining area.Table 1
Results of geostress measurement.
No.
Monitoring station
Depth (m)
Vertical principal stress (σv/MPa)
Maximum horizontal principal stress (σH/MPa)
Minimum horizontal principal stress (σH/MPa)
Maximum horizontal principal stress direction
1
First station
1040
27.67
51.17
27.11
N22.3°E
2
Second station
964
25.55
49.57
27.59
N31.2°E
3
Third station
967
25.64
49.74
26.15
N25.5°E
4
Fourth station
982
26.10
50.9
26.88
N20.1°E
5
Fifth station
1034
27.30
49.4
26.75
N35.7°ETable1 shows that geostress is mainly controlled by horizontal stress and that direction of the maximum principal stress is from N20.1 E to N35.7 E, which has consistency. The maximum horizontal principal stress is 1.81–1.95 times larger than the vertical stress, while the minimum horizontal principal stress is 0.98–1.03 times larger than the vertical stress. This can provide horizontal stress support for roadway horizontal deformation.


### 2.2. Roadway-Surrounding Rock Characteristics and Support

#### 2.2.1. Roadway-Surrounding Rock Characteristics

According to the coal lithology exposed by drilling holes in the mining area, direct roof of the No. 3 coal is 8.5 m medium sandstone, which is gray and of poor sorting. The main roof is 16.2 m fine sandstone, with gray-white and wave bedding. The direct bottom is 2.8 m mudstone that is rich in plant fossils. The main bottom is 3.2 m medium-fine sandstone, which is light gray and contains mica and cuttings. Figure3 lists the rock properties of the No. 3 coal seam roof and floor.Figure 3
No. 3 coal seam roof and floor lithology.


#### 2.2.2. Roadway Support Parameters

The No. 3401 working face roadway haulage is a rectangular section. The roadway section width is 5.5 m, and the roadway height is 4.0 m. Table2 lists the roadway support parameters, while Figure 4 shows the roadway support section.Table 2
Roadway support parameters.
Interval distance (mm)
Length (mm)
Diameter (mm)
Roof anchor (cable)/root number
Roadway-side anchor/root number
Anchor (cable) material
Remarks
Anchor
1000 × 1000
2400
22
6
4
Thread steel
1 ∗ 19 core steel strand is used for anchor cable
Anchor cable
2000 × 2300
7200
21.6
3
0
Steel strandFigure 4
Layout of roadway support section.


### 2.3. Roadway Floor Heave

To monitor microseismic and rock burst events during the excavation, the high-precision SOS microseismic monitoring system was adopted in the mine. Figure5 shows the structure of monitoring system, and Figure 6 shows the microseismic energy distribution of roadway during roadway haulage excavation.Figure 5
Microseismic monitoring system structure.Figure 6
Microseismic energy distribution.Figure6 shows that during August 4 to 12, the early excavation stage for roadway was influenced by horizontal tectonic stress and excavation disturbance. Microseismic events occurred frequently, but relatively little energy was released; this was considered the energy accumulation period. During August 13 to 29, microseismic events were in the active stage and the significantly increased energy was released, with the roadway constantly advancing in the anticlinal axis. During August 30 to September 8, the energy released by microseismic events increased further and the roadway started exhibiting dynamic stress phenomenon when excavation reached 160 m. Three roadway floor rock burst events were recorded during this period. However, the roadway roof and sidewalls had no obvious damage and maintained their integrity. Figure 7 shows the deformation and destruction of roadway-surrounding rock. These findings indicated the significance of studying roadway floor rock burst induced by mining disturbance, such as excavation disturbance and roof movement, to ensure mine safety.Figure 7
Deformation and failure of roadway-surrounding rock. (a) Roof. (b) Roadway side. (c) Floor. (d) Floor.
(a)
(b)
(c)
(d)


## 2.1. Engineering Geological Characteristics of Working Face

The No. 3401 working face is the first working face in the No. 4 mining area in Shandong Mine, China, and the mining seam belongs to No. 3 coal seam. The working face elevation is −986 m to −914 m, corresponding to the ground elevation of +52.86 m to +54.75 m, and the average buried depth of working face is 1000 m, which is classified as a deep mining face. The average thickness of coal seam is 4.7 m, and it is affected by an anticlinal structure. The working face layout is shown in Figure1. Based on the 15 boreholes in the mine fine-searching stage, macroscopic structure of the No. 3 coal seam is plotted in Figure 2. These boreholes only have boreholes No. T9-6 and No. T10-3 passing through No. 3401 working face.Figure 1
Layout of No. 3401 working face.Figure 2
Macroscopic storage structure profile of No. 3 coal seam.Figures1 and 2 show that overall occurrence status of the No. 3 coal mine is controlled by the fold structure. The anticlinal structure greatly influences the No. 3 coal seam uplift, while the influence of synclinal structure is relatively small. The coal seam in the No. 3401 working face is greatly influenced by the anticline, and the geostress is large. Therefore, roadway excavation activity further increases the possibility of rock burst. Table 1 shows the results of a geostress test in the mining area.Table 1
Results of geostress measurement.
No.
Monitoring station
Depth (m)
Vertical principal stress (σv/MPa)
Maximum horizontal principal stress (σH/MPa)
Minimum horizontal principal stress (σH/MPa)
Maximum horizontal principal stress direction
1
First station
1040
27.67
51.17
27.11
N22.3°E
2
Second station
964
25.55
49.57
27.59
N31.2°E
3
Third station
967
25.64
49.74
26.15
N25.5°E
4
Fourth station
982
26.10
50.9
26.88
N20.1°E
5
Fifth station
1034
27.30
49.4
26.75
N35.7°ETable1 shows that geostress is mainly controlled by horizontal stress and that direction of the maximum principal stress is from N20.1 E to N35.7 E, which has consistency. The maximum horizontal principal stress is 1.81–1.95 times larger than the vertical stress, while the minimum horizontal principal stress is 0.98–1.03 times larger than the vertical stress. This can provide horizontal stress support for roadway horizontal deformation.


## 2.2. Roadway-Surrounding Rock Characteristics and Support

### 2.2.1. Roadway-Surrounding Rock Characteristics

According to the coal lithology exposed by drilling holes in the mining area, direct roof of the No. 3 coal is 8.5 m medium sandstone, which is gray and of poor sorting. The main roof is 16.2 m fine sandstone, with gray-white and wave bedding. The direct bottom is 2.8 m mudstone that is rich in plant fossils. The main bottom is 3.2 m medium-fine sandstone, which is light gray and contains mica and cuttings. Figure3 lists the rock properties of the No. 3 coal seam roof and floor.Figure 3
No. 3 coal seam roof and floor lithology.


### 2.2.2. Roadway Support Parameters

The No. 3401 working face roadway haulage is a rectangular section. The roadway section width is 5.5 m, and the roadway height is 4.0 m. Table2 lists the roadway support parameters, while Figure 4 shows the roadway support section.Table 2
Roadway support parameters.
Interval distance (mm)
Length (mm)
Diameter (mm)
Roof anchor (cable)/root number
Roadway-side anchor/root number
Anchor (cable) material
Remarks
Anchor
1000 × 1000
2400
22
6
4
Thread steel
1 ∗ 19 core steel strand is used for anchor cable
Anchor cable
2000 × 2300
7200
21.6
3
0
Steel strandFigure 4
Layout of roadway support section.


## 2.2.1. Roadway-Surrounding Rock Characteristics

According to the coal lithology exposed by drilling holes in the mining area, direct roof of the No. 3 coal is 8.5 m medium sandstone, which is gray and of poor sorting. The main roof is 16.2 m fine sandstone, with gray-white and wave bedding. The direct bottom is 2.8 m mudstone that is rich in plant fossils. The main bottom is 3.2 m medium-fine sandstone, which is light gray and contains mica and cuttings. Figure3 lists the rock properties of the No. 3 coal seam roof and floor.Figure 3
No. 3 coal seam roof and floor lithology.


## 2.2.2. Roadway Support Parameters

The No. 3401 working face roadway haulage is a rectangular section. The roadway section width is 5.5 m, and the roadway height is 4.0 m. Table2 lists the roadway support parameters, while Figure 4 shows the roadway support section.Table 2
Roadway support parameters.
Interval distance (mm)
Length (mm)
Diameter (mm)
Roof anchor (cable)/root number
Roadway-side anchor/root number
Anchor (cable) material
Remarks
Anchor
1000 × 1000
2400
22
6
4
Thread steel
1 ∗ 19 core steel strand is used for anchor cable
Anchor cable
2000 × 2300
7200
21.6
3
0
Steel strandFigure 4
Layout of roadway support section.


## 2.3. Roadway Floor Heave

To monitor microseismic and rock burst events during the excavation, the high-precision SOS microseismic monitoring system was adopted in the mine. Figure5 shows the structure of monitoring system, and Figure 6 shows the microseismic energy distribution of roadway during roadway haulage excavation.Figure 5
Microseismic monitoring system structure.Figure 6
Microseismic energy distribution.Figure6 shows that during August 4 to 12, the early excavation stage for roadway was influenced by horizontal tectonic stress and excavation disturbance. Microseismic events occurred frequently, but relatively little energy was released; this was considered the energy accumulation period. During August 13 to 29, microseismic events were in the active stage and the significantly increased energy was released, with the roadway constantly advancing in the anticlinal axis. During August 30 to September 8, the energy released by microseismic events increased further and the roadway started exhibiting dynamic stress phenomenon when excavation reached 160 m. Three roadway floor rock burst events were recorded during this period. However, the roadway roof and sidewalls had no obvious damage and maintained their integrity. Figure 7 shows the deformation and destruction of roadway-surrounding rock. These findings indicated the significance of studying roadway floor rock burst induced by mining disturbance, such as excavation disturbance and roof movement, to ensure mine safety.Figure 7
Deformation and failure of roadway-surrounding rock. (a) Roof. (b) Roadway side. (c) Floor. (d) Floor.
(a)
(b)
(c)
(d)


## 3. Theoretical Analysis of Roadway Floor Stability

After excavation, deformation and damage of the roadway floor is mainly caused by horizontal stress and floor rock burst can easily occur [10]. The roadway floor’s width is recorded as L with the unit length to analyze the stress, as shown in Figure 8(a). The stress analysis of floor deformation is shown in Figure 8(b).Figure 8
Simplified model of the roadway floor force. (a) Stress analysis of roadway-surrounding rock. (b) Stress analysis of floor deformation.
(a)
(b)


### 3.1. Criterion Analysis of Roadway Floor Deformation

Based on the theory of material mechanics, the formulas for floor rock beam deflection and energy are as follows:(1)d2wdx2=MxEI=FωEI+M0EI,(2)EId2wdx2=M,(3)dVε=M2x2EIdx,where ω is the deformation of roadway floor, F is the roadway floor horizontal force, E is the elastic modulus of floor rock, I is the moment of inertia of floor rock, EI is the rock beam stiffness, M0 is the rock beam moment, and dVε is the unit rock beam energy.k is introduced to simplify the calculation:(4)k2=FEI.Equation (4) is substituted into equation (1):(5)d2wdx2+k2ω=M0EI.A general solution can be derived for equation (5):(6)ω=Asinkx+coskx+M0F,where A and B are the integral constant terms of equation (5).Using the stress analysis results for the floor rock beam, the initial condition of the beam can be deduced:(7)ω0=0,dωdxx=0=0,dωdxx=l/2=0,ωl=0,dωdxx=l=0,where l is the roadway width.Equation (7) is substituted into equation (6), and deflection is deduced as shown in the following equation:(8)ω=M0F1−coskx,where k=2nπ/l, (n = 1, 2, 3, ……), is substituted into equation (4). The floor horizontal force is obtained as shown in the following equation:(9)F=4n2π2EIl2,where n∈N+ indicates that the horizontal force F for maintaining floor rock stability beam has multiple values. Therefore, when n = 1, the horizontal force F reaches a minimum. That is, the minimum force Fcr is the critical stability force for the floor rock beam, as shown in the following equation:(10)Fcr=4π2EIl2.The instability of the roadway floor rock beamF>Fcr can be determined using equation (10). The roadway floor has undergone bending deformation and even damage, thus deteriorating the safety and stability of the roadway.


### 3.2. Energy Analysis of the Roadway Deformation and Instability

In the process of roadway excavation, the laws of deformation and failure for surrounding rock are affected by mining dynamic disturbances such as blasting, roof movement, and adjacent working face mining, which can be distinguished from the surrounding rock under static loading.Based on the law of thermodynamics, energy conversion is the essence of the physical change of the material. The deformation and failure of roadway floor can be attributed to four processes, namely, energy input, energy accumulation, energy dissipation, and energy release. In this article, the roadway floor rock burst is preliminarily discussed based on the energy principle, to provide a reference for other researchers.According to the attenuation law for dynamic stress wave propagated in rock [34], the energy input ED of the dynamic stress wave to the roadway floor in the process of surrounding rock propagation is shown in the following equation:(11)ED=ED0−EB=ED0r−r0−η,where ED0 is the initial energy value at the earthquake center, and it can be obtained by using a microseismic monitoring system. EB is the energy attenuated by stress wave in the process of propagating on the roadway floor. η is the stress wave attenuation coefficient, which is determined from the properties of rock mass. r and r0 are, respectively, expressed as the radius of roadway and the earthquake source distance from the roadway center.Equations (1)–(3) and (6) are combined, and accumulated strain energy expression for the microsection of the rock beam is as follows:(12)dVε=4n4π4M02EIF2l41+cos4nπlxdx.Based on the minimum energy principle of rock failure, energy released from the roadway-surrounding rock is bigger than the energy consumed in a rock burst [35]. The energy criterion for rock impact instability can be derived as follows:(13)Ec=σc22E,orEc=τc22E,(14)E0=∫dVε+ED−Ec>0,where E0 is the residual elastic energy, ED is the roadway-surrounding energy, and EC is the energy consumed by rock mass destruction. σc and τc are, respectively, expressed as the uniaxial compressive strength and shear strength, while E is the rock mass elastic modulus.Equations (11)–(14) are combined, and the energy criterion for a rock burst is deduced as shown in the following equation:(15)E0=4n4π4M02EIF2l4∫1+cos4nπlxdx+ED0r−r0−η−σc22E>0.According to equation (15), the residual energy E0 of roadway floor is closely related to the bending moment M and its location, which are influenced by the roadway buried depth h, roadway span l, horizontal force F, and rock stiffness EI. When the roadway span l, rock stiffness EI, elastic modulus of rock E, and Poisson’s ratio υ are constant, the greater the roadway buried depth and dynamic load strength, the greater the residual elastic energy of the roadway floor, accompanied by the development and penetration of fractures. If the elastic energy is released instantaneously, rock burst can occur in the roadway floor, which will endanger safe and efficient mine production.


## 3.1. Criterion Analysis of Roadway Floor Deformation

Based on the theory of material mechanics, the formulas for floor rock beam deflection and energy are as follows:(1)d2wdx2=MxEI=FωEI+M0EI,(2)EId2wdx2=M,(3)dVε=M2x2EIdx,where ω is the deformation of roadway floor, F is the roadway floor horizontal force, E is the elastic modulus of floor rock, I is the moment of inertia of floor rock, EI is the rock beam stiffness, M0 is the rock beam moment, and dVε is the unit rock beam energy.k is introduced to simplify the calculation:(4)k2=FEI.Equation (4) is substituted into equation (1):(5)d2wdx2+k2ω=M0EI.A general solution can be derived for equation (5):(6)ω=Asinkx+coskx+M0F,where A and B are the integral constant terms of equation (5).Using the stress analysis results for the floor rock beam, the initial condition of the beam can be deduced:(7)ω0=0,dωdxx=0=0,dωdxx=l/2=0,ωl=0,dωdxx=l=0,where l is the roadway width.Equation (7) is substituted into equation (6), and deflection is deduced as shown in the following equation:(8)ω=M0F1−coskx,where k=2nπ/l, (n = 1, 2, 3, ……), is substituted into equation (4). The floor horizontal force is obtained as shown in the following equation:(9)F=4n2π2EIl2,where n∈N+ indicates that the horizontal force F for maintaining floor rock stability beam has multiple values. Therefore, when n = 1, the horizontal force F reaches a minimum. That is, the minimum force Fcr is the critical stability force for the floor rock beam, as shown in the following equation:(10)Fcr=4π2EIl2.The instability of the roadway floor rock beamF>Fcr can be determined using equation (10). The roadway floor has undergone bending deformation and even damage, thus deteriorating the safety and stability of the roadway.


## 3.2. Energy Analysis of the Roadway Deformation and Instability

In the process of roadway excavation, the laws of deformation and failure for surrounding rock are affected by mining dynamic disturbances such as blasting, roof movement, and adjacent working face mining, which can be distinguished from the surrounding rock under static loading.Based on the law of thermodynamics, energy conversion is the essence of the physical change of the material. The deformation and failure of roadway floor can be attributed to four processes, namely, energy input, energy accumulation, energy dissipation, and energy release. In this article, the roadway floor rock burst is preliminarily discussed based on the energy principle, to provide a reference for other researchers.According to the attenuation law for dynamic stress wave propagated in rock [34], the energy input ED of the dynamic stress wave to the roadway floor in the process of surrounding rock propagation is shown in the following equation:(11)ED=ED0−EB=ED0r−r0−η,where ED0 is the initial energy value at the earthquake center, and it can be obtained by using a microseismic monitoring system. EB is the energy attenuated by stress wave in the process of propagating on the roadway floor. η is the stress wave attenuation coefficient, which is determined from the properties of rock mass. r and r0 are, respectively, expressed as the radius of roadway and the earthquake source distance from the roadway center.Equations (1)–(3) and (6) are combined, and accumulated strain energy expression for the microsection of the rock beam is as follows:(12)dVε=4n4π4M02EIF2l41+cos4nπlxdx.Based on the minimum energy principle of rock failure, energy released from the roadway-surrounding rock is bigger than the energy consumed in a rock burst [35]. The energy criterion for rock impact instability can be derived as follows:(13)Ec=σc22E,orEc=τc22E,(14)E0=∫dVε+ED−Ec>0,where E0 is the residual elastic energy, ED is the roadway-surrounding energy, and EC is the energy consumed by rock mass destruction. σc and τc are, respectively, expressed as the uniaxial compressive strength and shear strength, while E is the rock mass elastic modulus.Equations (11)–(14) are combined, and the energy criterion for a rock burst is deduced as shown in the following equation:(15)E0=4n4π4M02EIF2l4∫1+cos4nπlxdx+ED0r−r0−η−σc22E>0.According to equation (15), the residual energy E0 of roadway floor is closely related to the bending moment M and its location, which are influenced by the roadway buried depth h, roadway span l, horizontal force F, and rock stiffness EI. When the roadway span l, rock stiffness EI, elastic modulus of rock E, and Poisson’s ratio υ are constant, the greater the roadway buried depth and dynamic load strength, the greater the residual elastic energy of the roadway floor, accompanied by the development and penetration of fractures. If the elastic energy is released instantaneously, rock burst can occur in the roadway floor, which will endanger safe and efficient mine production.


## 4. Analysis of Failure Characteristics and Causes of Rock Burst

### 4.1. Failure Characteristics of Roadway Floor Rock Burst

#### 4.1.1. Borehole Detection Recorder Parameters

In light of the roadway haulage floor heave, the YTJ20 borehole detection recorder was used to detect cracks and separation in the roadway-surrounding rock. The recorder uses a high-resolution color peeposcope and LCD display. It can work continuously for 12 hours, achieving real-time image display and recording and storage with small volume, light weight, and simple operation. The borehole detection recorder and its main parameters are as shown in Table3 and Figure 9.Table 3
Main parameters of the YTJ20 borehole detection recorder.
Camera diameter (mm)
Camera length (mm)
Depth of detection (m)
Host size (length × width × height) (mm3)
Working hours (h)
Storage capacity (G)
25
100
20
240 × 190 × 83
8
2Figure 9
YTJ20 borehole detection recorder. (a) Detection recorder host. (b) Guide rod.
(a)
(b)


#### 4.1.2. Borehole Detection Monitoring Scheme

According to construction situation for the roadway, three detection sections were set at a specific distance from the roadway heading face at 100 m, 150 m, and 200 m, labeled Sections A, B, and C. Four detection boreholes were arranged in each detection section: roof drilling, roadway sidewalls drilling, and floor drilling. The “+” detection method was used for the borehole detection layout, with 12 monitoring boreholes in total. Borehole detection section layout is as shown in Figure10(a). The borehole diameter was 32 mm, and the drilling length was 7 m; each detection boreholes were arranged, as shown in Figure 10(b).Figure 10
Schematic diagram of detection section site layout.
(a)
(b)


#### 4.1.3. Analysis of the Detection Results

The borehole detection recorder was used to detect the rock surrounding each detection section. At the same time, the section 80 m away from the roadway head was taken as the base level for avoiding surrounding rock rupture, on August 11, 2016. The roadway excavation speed was 10 m/d, and sequence of borehole detection is shown in Table4.Table 4
Detection sequence of roadway boreholes.
Time
Section number
August 21, 2016
A section
August 26, 2016
B Section
August 31, 2016
C SectionAfter the detection was completed, video capture of the roadway roof, sidewalls, and floor was extracted as shown in Figures11–13.Figure 11
Video capture of the roof detection borehole. (a) Intact roof strata. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)Figure 12
Video capture of the roadway side detection borehole. (a) Complete coal seam. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)Figure 13
Video capture of the roadway floor detection borehole. (a) Complete floor rock stratum. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)By using the borehole recorder to detect destruction of roadway-surrounding rock, fracture distribution of the surrounding rock was analyzed. An ultrasound transmitter was used to detect failure destruction in the surrounding rock between adjacent boreholes. Main parameters of mine ultrasonic transmitter are shown in Table5 and Figure 14. The transmitter-monitoring point was identical with the borehole point. Firstly, borehole detection was carried out. Secondly, transmitter detection was carried out without rearranging the measuring points. Finally, rupture of surrounding rock occurred along the borehole from the outside to inside. The fracture area between adjacent boreholes was connected by using a spline curve, and the area between fracture areas was regarded as the complete area, as shown in Figure 15.Table 5
Main parameters of the mine ultrasonic transmitter.
Host size (length × width × height) (mm3)
Channel number
Ultrasonic time accuracy (με)
Working hours (h)
Transmitting voltage (V)
Storage capacity (G)
270 × 220 × 75
4
0.05
>5
65,250,500,1000
4Figure 14
Mine ultrasonic transmitter.Figure 15
Fracture distribution of the roadway-surrounding rock. (a) A Section. (b) B Section. (c) C Section.
(a)
(b)
(c)Figure15 shows that the fracture area of roadway-surrounding rock gradually narrowed from its width along the borehole axis direction. The closer the detection section is to the anticline axis, the more serious the fracture in the surrounding rock. Sections A, B, and C in the roadway can be divided into two, two, and three fracture areas, respectively. Roadway-surrounding rock damage was most serious in the range of 1 m-2 m. The damage was mainly breakage and fracturing, and the scope was called the loose area. The second fracture circle was 3 m–4.5 m and mainly consisted of cracks and fissures. Along the borehole, the third fracture circle was 5.5 m–6 m. Sections A and B had only a few cracks in the roadway-surrounding rock. In addition, the fracture distribution for roadway floor was more serious than the distribution for roadway roof and sidewalls based on comparison of the roadway-surrounding fracture.


### 4.2. Cause Analysis of Roadway Floor Rock Burst

For causes of roadway floor failure, first, from roadway rock lithology perspective, the roadway roof is 24.7 m thick sand, while the floor is 2.8 m sandy mudstone with low bearing capacity. Second, from the perspective of roadway supporting measures, the roadway roof and sidewalls were supported by anchor, anchor cable, and wire mesh. However, the floor was unsupported, which resulted in low floor resistance to deformation and damage. Finally, based on the stress state of roadway-surrounding rock, the roadway was excavated in the anticlinal area, which made the surrounding rock change from three-dimensional stress state to two-dimensional stress state. The roadway roof and sidewalls were supported and retained the three-dimensional stress state, while the roadway floor had a two-dimensional stress state without support. Furthermore, roadway-surrounding rock was influenced by high tectonic stress and excavation disturbance stress, which further aggravated the floor failure.


## 4.1. Failure Characteristics of Roadway Floor Rock Burst

### 4.1.1. Borehole Detection Recorder Parameters

In light of the roadway haulage floor heave, the YTJ20 borehole detection recorder was used to detect cracks and separation in the roadway-surrounding rock. The recorder uses a high-resolution color peeposcope and LCD display. It can work continuously for 12 hours, achieving real-time image display and recording and storage with small volume, light weight, and simple operation. The borehole detection recorder and its main parameters are as shown in Table3 and Figure 9.Table 3
Main parameters of the YTJ20 borehole detection recorder.
Camera diameter (mm)
Camera length (mm)
Depth of detection (m)
Host size (length × width × height) (mm3)
Working hours (h)
Storage capacity (G)
25
100
20
240 × 190 × 83
8
2Figure 9
YTJ20 borehole detection recorder. (a) Detection recorder host. (b) Guide rod.
(a)
(b)


### 4.1.2. Borehole Detection Monitoring Scheme

According to construction situation for the roadway, three detection sections were set at a specific distance from the roadway heading face at 100 m, 150 m, and 200 m, labeled Sections A, B, and C. Four detection boreholes were arranged in each detection section: roof drilling, roadway sidewalls drilling, and floor drilling. The “+” detection method was used for the borehole detection layout, with 12 monitoring boreholes in total. Borehole detection section layout is as shown in Figure10(a). The borehole diameter was 32 mm, and the drilling length was 7 m; each detection boreholes were arranged, as shown in Figure 10(b).Figure 10
Schematic diagram of detection section site layout.
(a)
(b)


### 4.1.3. Analysis of the Detection Results

The borehole detection recorder was used to detect the rock surrounding each detection section. At the same time, the section 80 m away from the roadway head was taken as the base level for avoiding surrounding rock rupture, on August 11, 2016. The roadway excavation speed was 10 m/d, and sequence of borehole detection is shown in Table4.Table 4
Detection sequence of roadway boreholes.
Time
Section number
August 21, 2016
A section
August 26, 2016
B Section
August 31, 2016
C SectionAfter the detection was completed, video capture of the roadway roof, sidewalls, and floor was extracted as shown in Figures11–13.Figure 11
Video capture of the roof detection borehole. (a) Intact roof strata. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)Figure 12
Video capture of the roadway side detection borehole. (a) Complete coal seam. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)Figure 13
Video capture of the roadway floor detection borehole. (a) Complete floor rock stratum. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)By using the borehole recorder to detect destruction of roadway-surrounding rock, fracture distribution of the surrounding rock was analyzed. An ultrasound transmitter was used to detect failure destruction in the surrounding rock between adjacent boreholes. Main parameters of mine ultrasonic transmitter are shown in Table5 and Figure 14. The transmitter-monitoring point was identical with the borehole point. Firstly, borehole detection was carried out. Secondly, transmitter detection was carried out without rearranging the measuring points. Finally, rupture of surrounding rock occurred along the borehole from the outside to inside. The fracture area between adjacent boreholes was connected by using a spline curve, and the area between fracture areas was regarded as the complete area, as shown in Figure 15.Table 5
Main parameters of the mine ultrasonic transmitter.
Host size (length × width × height) (mm3)
Channel number
Ultrasonic time accuracy (με)
Working hours (h)
Transmitting voltage (V)
Storage capacity (G)
270 × 220 × 75
4
0.05
>5
65,250,500,1000
4Figure 14
Mine ultrasonic transmitter.Figure 15
Fracture distribution of the roadway-surrounding rock. (a) A Section. (b) B Section. (c) C Section.
(a)
(b)
(c)Figure15 shows that the fracture area of roadway-surrounding rock gradually narrowed from its width along the borehole axis direction. The closer the detection section is to the anticline axis, the more serious the fracture in the surrounding rock. Sections A, B, and C in the roadway can be divided into two, two, and three fracture areas, respectively. Roadway-surrounding rock damage was most serious in the range of 1 m-2 m. The damage was mainly breakage and fracturing, and the scope was called the loose area. The second fracture circle was 3 m–4.5 m and mainly consisted of cracks and fissures. Along the borehole, the third fracture circle was 5.5 m–6 m. Sections A and B had only a few cracks in the roadway-surrounding rock. In addition, the fracture distribution for roadway floor was more serious than the distribution for roadway roof and sidewalls based on comparison of the roadway-surrounding fracture.


## 4.1.1. Borehole Detection Recorder Parameters

In light of the roadway haulage floor heave, the YTJ20 borehole detection recorder was used to detect cracks and separation in the roadway-surrounding rock. The recorder uses a high-resolution color peeposcope and LCD display. It can work continuously for 12 hours, achieving real-time image display and recording and storage with small volume, light weight, and simple operation. The borehole detection recorder and its main parameters are as shown in Table3 and Figure 9.Table 3
Main parameters of the YTJ20 borehole detection recorder.
Camera diameter (mm)
Camera length (mm)
Depth of detection (m)
Host size (length × width × height) (mm3)
Working hours (h)
Storage capacity (G)
25
100
20
240 × 190 × 83
8
2Figure 9
YTJ20 borehole detection recorder. (a) Detection recorder host. (b) Guide rod.
(a)
(b)


## 4.1.2. Borehole Detection Monitoring Scheme

According to construction situation for the roadway, three detection sections were set at a specific distance from the roadway heading face at 100 m, 150 m, and 200 m, labeled Sections A, B, and C. Four detection boreholes were arranged in each detection section: roof drilling, roadway sidewalls drilling, and floor drilling. The “+” detection method was used for the borehole detection layout, with 12 monitoring boreholes in total. Borehole detection section layout is as shown in Figure10(a). The borehole diameter was 32 mm, and the drilling length was 7 m; each detection boreholes were arranged, as shown in Figure 10(b).Figure 10
Schematic diagram of detection section site layout.
(a)
(b)


## 4.1.3. Analysis of the Detection Results

The borehole detection recorder was used to detect the rock surrounding each detection section. At the same time, the section 80 m away from the roadway head was taken as the base level for avoiding surrounding rock rupture, on August 11, 2016. The roadway excavation speed was 10 m/d, and sequence of borehole detection is shown in Table4.Table 4
Detection sequence of roadway boreholes.
Time
Section number
August 21, 2016
A section
August 26, 2016
B Section
August 31, 2016
C SectionAfter the detection was completed, video capture of the roadway roof, sidewalls, and floor was extracted as shown in Figures11–13.Figure 11
Video capture of the roof detection borehole. (a) Intact roof strata. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)Figure 12
Video capture of the roadway side detection borehole. (a) Complete coal seam. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)Figure 13
Video capture of the roadway floor detection borehole. (a) Complete floor rock stratum. (b) Fracture. (c) Crack. (d) Broken.
(a)
(b)
(c)
(d)By using the borehole recorder to detect destruction of roadway-surrounding rock, fracture distribution of the surrounding rock was analyzed. An ultrasound transmitter was used to detect failure destruction in the surrounding rock between adjacent boreholes. Main parameters of mine ultrasonic transmitter are shown in Table5 and Figure 14. The transmitter-monitoring point was identical with the borehole point. Firstly, borehole detection was carried out. Secondly, transmitter detection was carried out without rearranging the measuring points. Finally, rupture of surrounding rock occurred along the borehole from the outside to inside. The fracture area between adjacent boreholes was connected by using a spline curve, and the area between fracture areas was regarded as the complete area, as shown in Figure 15.Table 5
Main parameters of the mine ultrasonic transmitter.
Host size (length × width × height) (mm3)
Channel number
Ultrasonic time accuracy (με)
Working hours (h)
Transmitting voltage (V)
Storage capacity (G)
270 × 220 × 75
4
0.05
>5
65,250,500,1000
4Figure 14
Mine ultrasonic transmitter.Figure 15
Fracture distribution of the roadway-surrounding rock. (a) A Section. (b) B Section. (c) C Section.
(a)
(b)
(c)Figure15 shows that the fracture area of roadway-surrounding rock gradually narrowed from its width along the borehole axis direction. The closer the detection section is to the anticline axis, the more serious the fracture in the surrounding rock. Sections A, B, and C in the roadway can be divided into two, two, and three fracture areas, respectively. Roadway-surrounding rock damage was most serious in the range of 1 m-2 m. The damage was mainly breakage and fracturing, and the scope was called the loose area. The second fracture circle was 3 m–4.5 m and mainly consisted of cracks and fissures. Along the borehole, the third fracture circle was 5.5 m–6 m. Sections A and B had only a few cracks in the roadway-surrounding rock. In addition, the fracture distribution for roadway floor was more serious than the distribution for roadway roof and sidewalls based on comparison of the roadway-surrounding fracture.


## 4.2. Cause Analysis of Roadway Floor Rock Burst

For causes of roadway floor failure, first, from roadway rock lithology perspective, the roadway roof is 24.7 m thick sand, while the floor is 2.8 m sandy mudstone with low bearing capacity. Second, from the perspective of roadway supporting measures, the roadway roof and sidewalls were supported by anchor, anchor cable, and wire mesh. However, the floor was unsupported, which resulted in low floor resistance to deformation and damage. Finally, based on the stress state of roadway-surrounding rock, the roadway was excavated in the anticlinal area, which made the surrounding rock change from three-dimensional stress state to two-dimensional stress state. The roadway roof and sidewalls were supported and retained the three-dimensional stress state, while the roadway floor had a two-dimensional stress state without support. Furthermore, roadway-surrounding rock was influenced by high tectonic stress and excavation disturbance stress, which further aggravated the floor failure.


## 5. Roadway Floor Rock Burst Control Technology

### 5.1. Design of Floor Rock Burst Control Technology

Based on the stability analysis and failure characteristics of roadway-surrounding rock, the integrity and stability of roadway roof and sidewalls can be maintained by using the original support parameters. Therefore, the roadway support design was only implemented for the roadway floor. The “overbreaking-bolting and grouting-backfill” method was put forward for controlling roadway floor rock burst, while the support parameters for roadway roof and sidewalls remain unchanged. The support parameters optimization processes for the roadway are shown in Figures16 and 17.Figure 16
Construction flowchart of the roadway floor.Figure 17
Roadway support parameter optimization design diagram. (a) Overbreaking. (b) Bolting and grouting (c) Backfill.
(a)
(b)
(c)The main support parameters for the roadway floor were as follows:(1)
Roadway floor overbreakRock mass from within 600 mm below the roadway floor was excavated, as shown in Figure17(a).(2)
Grouting boltAfter excavation, the left-lateral and high-strength grouting bolts were used to support the roadway floor. The specification of bolt was Φ25 mm × 2,400 mm, and the bolt spacing was 1,000 mm, corresponding to the position of roof bolts, as shown in Figure17(b). The grouting pressure was 2.3 MPa–3.4 MPa.(3)
Concrete groutingAfter grouting bolt installation, the roadway floor was grouted with the thickness 20 mm of spray layer, for maintaining smoothness and integrity of the floor, and the grouting port of bolt should be protected to prevent blockage.(4)
Grouting slurryThe PO42.5 alumino-silicate cement single slurry was selected as the grouting material, and the water-cement ratio was 1 : 2.5.(5)
Concrete backfillingConcrete of strength grade C40 was used to backfill the overbreak portion of roadway floor, as shown in Figure17(c).In addition, according to roadway’s actual excavation situation for the roadway, reinforcement of the roadway floor was carried out 50–70 m behind the roadway working face.According to the above steps, improved floor lithology can effectively alleviate the floor’s stress effect. The cracked rock can be cemented by grouting, and the integral bearing capacity of roadway floor will thus be enhanced.


### 5.2. Field Application

Deformation of the roadway in the test section was monitored for three months to evaluate the support effect of the roadway floor optimization scheme. The monitoring content included roof subsidence, floor heave, and displacement of the roadway sidewalls. The measuring point of roadway-surrounding rock was arranged using the “+” measuring method. Monitoring pointsA and C were located at 1.375 m from the roadway floor to roadway side. Monitoring point B was located at 2.75 from the floor to roadway side. Monitoring points D and D1 were located at 2.0 m from the roadway side to floor, as shown in Figure 18.Figure 18
Monitoring point layout for the roadway surface.Based on monitoring deformation data for the roadway-surrounding rock, deformation distribution of the roadway-surrounding rock was plotted as shown in Figure19.Figure 19
Displacement distribution curves of roadway-surrounding rock. (a) Roadway sides. (b) Roof. (c) Floor.
(a)
(b)
(c)Figure19 showed that the deformation of roadway-surrounding rock entered into an active phase after excavation. Roof subsidence, convergence deformation of the roadway sidewalls, and floor heave were more obvious. When the distance between reinforcing position of the floor and roadway working face was 50–70 m, overbreaking, grouting, and backfilling started occurring, and the lithology of roadway floor improved with roadway excavation. After grouted support of the roadway floor, the floor’s deformation rate gradually changed from 1.3 mm/d to 0.33 mm/d. The roadway floor deformation changed from the rapid deformation phase to slow deformation phase, and the deformation eventually stabilized. The floor lithology also had a better control effect on the deformation of the roadway roof and sidewalls, thus guaranteeing stability and safety of the roadway-surrounding rock. The support effect of roadway after optimizing the support parameter is shown in Figure 20. The distribution of roadway microseismic energy during excavation ifs shown in Figure 21.Figure 20
Roadway support parameters optimization effect.Figure 21
Distribution of roadway microseismic energy.


## 5.1. Design of Floor Rock Burst Control Technology

Based on the stability analysis and failure characteristics of roadway-surrounding rock, the integrity and stability of roadway roof and sidewalls can be maintained by using the original support parameters. Therefore, the roadway support design was only implemented for the roadway floor. The “overbreaking-bolting and grouting-backfill” method was put forward for controlling roadway floor rock burst, while the support parameters for roadway roof and sidewalls remain unchanged. The support parameters optimization processes for the roadway are shown in Figures16 and 17.Figure 16
Construction flowchart of the roadway floor.Figure 17
Roadway support parameter optimization design diagram. (a) Overbreaking. (b) Bolting and grouting (c) Backfill.
(a)
(b)
(c)The main support parameters for the roadway floor were as follows:(1)
Roadway floor overbreakRock mass from within 600 mm below the roadway floor was excavated, as shown in Figure17(a).(2)
Grouting boltAfter excavation, the left-lateral and high-strength grouting bolts were used to support the roadway floor. The specification of bolt was Φ25 mm × 2,400 mm, and the bolt spacing was 1,000 mm, corresponding to the position of roof bolts, as shown in Figure17(b). The grouting pressure was 2.3 MPa–3.4 MPa.(3)
Concrete groutingAfter grouting bolt installation, the roadway floor was grouted with the thickness 20 mm of spray layer, for maintaining smoothness and integrity of the floor, and the grouting port of bolt should be protected to prevent blockage.(4)
Grouting slurryThe PO42.5 alumino-silicate cement single slurry was selected as the grouting material, and the water-cement ratio was 1 : 2.5.(5)
Concrete backfillingConcrete of strength grade C40 was used to backfill the overbreak portion of roadway floor, as shown in Figure17(c).In addition, according to roadway’s actual excavation situation for the roadway, reinforcement of the roadway floor was carried out 50–70 m behind the roadway working face.According to the above steps, improved floor lithology can effectively alleviate the floor’s stress effect. The cracked rock can be cemented by grouting, and the integral bearing capacity of roadway floor will thus be enhanced.


## 5.2. Field Application

Deformation of the roadway in the test section was monitored for three months to evaluate the support effect of the roadway floor optimization scheme. The monitoring content included roof subsidence, floor heave, and displacement of the roadway sidewalls. The measuring point of roadway-surrounding rock was arranged using the “+” measuring method. Monitoring pointsA and C were located at 1.375 m from the roadway floor to roadway side. Monitoring point B was located at 2.75 from the floor to roadway side. Monitoring points D and D1 were located at 2.0 m from the roadway side to floor, as shown in Figure 18.Figure 18
Monitoring point layout for the roadway surface.Based on monitoring deformation data for the roadway-surrounding rock, deformation distribution of the roadway-surrounding rock was plotted as shown in Figure19.Figure 19
Displacement distribution curves of roadway-surrounding rock. (a) Roadway sides. (b) Roof. (c) Floor.
(a)
(b)
(c)Figure19 showed that the deformation of roadway-surrounding rock entered into an active phase after excavation. Roof subsidence, convergence deformation of the roadway sidewalls, and floor heave were more obvious. When the distance between reinforcing position of the floor and roadway working face was 50–70 m, overbreaking, grouting, and backfilling started occurring, and the lithology of roadway floor improved with roadway excavation. After grouted support of the roadway floor, the floor’s deformation rate gradually changed from 1.3 mm/d to 0.33 mm/d. The roadway floor deformation changed from the rapid deformation phase to slow deformation phase, and the deformation eventually stabilized. The floor lithology also had a better control effect on the deformation of the roadway roof and sidewalls, thus guaranteeing stability and safety of the roadway-surrounding rock. The support effect of roadway after optimizing the support parameter is shown in Figure 20. The distribution of roadway microseismic energy during excavation ifs shown in Figure 21.Figure 20
Roadway support parameters optimization effect.Figure 21
Distribution of roadway microseismic energy.


## 6. Results

This article analyzed the deformation and failure characteristics of roadway-surrounding rock based on the roadway floor deformation induced by floor rock burst. Then, the “overbreaking-bolting and grouting-backfill” method was put forward to prevent roadway floor rock burst and heave and to guarantee roadway stability. The main results obtained were as follows:(1)
Deformation instability criterion for the roadway floor was deduced using the theory of material mechanics. The energy transformation of floor rock burst was discussed preliminarily using the energy conservation principle. We concluded that the greater the depth and dynamic load of the roadway, the greater the floor elastic energy storage, thus increasing the likelihood of a roadway floor rock burst.(2)
Using a borehole recorder to determine the internal characteristics of roadway-surrounding rock, the fractured area in the roadway-surrounding rock was found to narrow gradually in terms of its width along the borehole axis direction. The closer the detection section was to the anticlinal axis, the more the serious fracturing of the surrounding rock. In addition, the roadway floor contained many cracks and broken areas, whereas cracks and fissures occurred in the roadway roof and sidewalls.(3)
Based on the deformation and failure characteristics of deep roadways, the “overbreaking-bolting and grouting-backfill” method was put forward and tested in the field. The results showed that the floor’s deformation was effectively controlled and the total deformation was small, thus ensuring the stability of the roadway-surrounding rock.



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*Source: 7938491-2019-08-21.xml*