Investigation on the Failure Mechanism of Weak Floors in Deep and High-Stress Roadway and the Corresponding Control Technology

Abstract

Large deformation of roadway and floor burst are the two major geotechnical hazards encountered with high mining stress in deep mines. In this paper, the stress and energy conditions generated by the impact damage on the rock surrounding a roadway are analyzed, and UDEC software was used to study the deformation characteristics of the roadway, as well as its failure mechanism under the influence of superimposed dynamic and static loads. The results indicate that the soft floor of a deep-buried roadway has a high damage degree and an obvious stress release effect, high static load leads to slow floor heave, and strong dynamic load disturbance is the principal trigger leading to floor burst. In addition, the anisotropy caused by the bedding surface weakens the cooperative characteristics of the support system, resulting in serious instability of the whole rock surrounding the roadway. Full-section anchor cables and inverted arches were adopted to maintain the stability of the rock surrounding the roadway. The monitoring results obtained from field tests show that the adoption of the combined support system effectively avoids floor burst caused by the superposition of dynamic and static loads; the maximum floor heave is 67.9 mm, which is 95% lower than the original value, ensuring safety in coal mining operations.

1. Introduction

The increasing demand and consumption of coal resources has made shallow coal mining very rare and nowadays, most mining activities are taking place at a deep level. Therefore, many geotechnical hazards, such as floor burst, rib shrinkage, and roof caving are inevitable in roadways, especially those with high mining-induced stress [1,2,3,4]. Due to the influence of traditional mining supports and the limitation of on-site production, the stability of the floors in roadways has always been neglected, and less support is applied for the maintenance of roadways. As a consequence, floor burst usually appears in roadways with strong dynamic pressure [5]. Floor burst is becoming more serious with the extension of coal mining depth [6,7,8,9], posing huge threats to safety in coal mines.

Generally, floor heave is a significant feature of potential floor burst [10] and many studies have already been conducted on the types of floor heave and control measures for them. For example, four types of floor heave (namely, extrusion flow type, flexure fold type, shear dislocation type, and water-induced swelling type) were documented in the research performed by Chen and Lu et al. [11]. Bai et al. [12] described the characteristics of stress distribution, and proposed to eliminate floor heave by strengthening the floor and ribs with full-length anchor bolts. In the study of Xu et al. [13], steel piles are applied to control the floor heave of gob-side entry retaining successfully. Zhao et al. [14] established the mechanical model of integral support and theoretically deduced the bearing capacity of the support and the critical instability load of the reverse arch. With the adoption of similar simulations and numerical calculations, Guo et al. [15] put forward an anti-floor heave technology by reinforcing the ribs and corners of roadways. Zhang et al. [16] managed to control floor heave by using grouting reinforcement.

In order to investigate the failure mechanism of floor burst in roadway, many efforts have been taken by mining engineers and scholars as well. Dou et al. [17] theoretically analyzed the energy and stress conditions of rock burst and then put forward the principle of floor burst induced by the superposition of dynamic and static loads. He and Dou et al. [18] proposed a way of rock burst assessment and prediction by dynamic-static stress. He et al. [19] proposed a new way to classify rock burst and developed a novel energy-absorbing bolt to resist both static and impact loadings. Cao et al. [20] proposed to decrease the bending strength of floors to reduce the possibility of the occurrence of floor burst. Gao et al. [21] established a strong-soft-strong structural model; the stress criterion and energy criterion of rock burst failure were analyzed as well. Moreover, Kang et al. [22] applied a combined “ground support-rock modification-destressing” strategy to the control of the large deformation of the roadway influenced by mining stress. Xiao [23] concluded that floor burst is mainly dominated by the high static load, but occurrence is induced by the strong dynamic load.

The abovementioned studies analyze the failure mechanism of floor burst and the corresponding control technology, which provides a theoretical and technical basis for the management of rock burst. However, for coal mines with different geological conditions, the controlling effect of floor heave varies, even with same support parameters [24,25,26].

Given the seriously damaged mudstone floor in the No. −830 roadway of Zhangshuanglou coal mine, this manuscript puts forward a combined support technology characterized by the usage of cables and combined arched girders. It aims to avoid floor burst by first adopting anchor cables to maintain the stability of the floor and ribs. Thereafter, a set of arched girders are used to further reinforce the floor. The application shows that the combined support technology has a significant effect on floor burst control in deep and high-stress roadway, which is beneficial to the improvement of mining productivity.

2. Engineering Geological Conditions

2.1. Geological Condition

Zhangshuanglou Coal Mine is located in Pei County, Xuzhou City, Jiangsu Province, China (as depicted in Figure 1a). At present, the main producing coal seams are the No. 7 coal seam and No. 9 coal seam, with a design production capacity of 1.2 million tons per year. The average thickness of the No. 7 coal seam is 1.4 m, while the dip angle ranges from 18° to 26°. The buried depth of the No. −830 roadway is 870 m, which is sited at the junction of mudstone and sandstone above the No. 7 coal seam. The position of the No. −830 roadway and the synthesis histogram around the coal seams are presented in Figure 1b. The width and length of the No. 7102 workface are 117 m and 680 m respectively. The advancing direction is approximately parallel to the strike of the No. −830 main roadway. Besides, there is a 30 m coal pillar exiting between the return air roadway of No. 7102 workface and the No. −830 main roadway (Figure 1c).

2.2. Supporting System and Deformation Characteristics

The section shape of the No. −830 roadway is a semi-circular arch, 5.2 m in width and 4.4 m in height, as shown in Figure 1d. Meanwhile, the height of the straight wall is 1.8 m and the radius of the arch is 2.6 m. From the figure, it can also be seen that both anchor bolts and cables are used in the initial support. The bolts are made with Q335 steel and have a diameter of 22 mm and a length of 2400 mm. The size of the pallets that go with the bolts is 120 mm × 120 mm × 10 mm. The rod and row spacing are both 800 mm, while the angle between the side bottom anchor bolts and the floor is 30°. In contrast, the specifications of the anchor cables used are 18.9 mm in diameter and 6250 in length, and the size of the cable pallets is 260 mm × 260 mm × 12 mm. The rod and row spacing are 1600 mm and 2250 mm, respectively.

During the advancement of the No. 7102 workface, high mining-induced stress appears in the No. −830 roadway. Consequently, the base of the conveyor belt was migrated to the middle of the roadway. Occasionally, the anchor bolts were even destroyed and roof caving took place due to the mining-induced high stress (Figure 1e). The maximum floor heave obtained from the field can reach 1241 mm and the maximum subsidence of the roof is up to 247 mm, which seriously restrict the ongoing mining activity.

2.3. Damage Determination of the Surrounding Rock Caused by Rock Burst

Damage determination is not only an important quantitative index to evaluate the quality of the adopted anchorage system, but also the theoretical basis for the determination of the supporting parameters. In order to further quantify the actual damage in the No. −830 roadway, several boreholes (with a depth of 4 m) were drilled in the floor and two side ribs. Thereafter, ZDY borehole peeps were set inside for quantitative assessment of the damage. The damage characteristics of the surrounding rock are shown in Figure 2.

(1) Borehole on right rib: horizontal and longitudinal cracks are observed at a depth of 3.6 m. When the depth decreases to 1.5 m, serious rock damage was noticed and local cavities appear. The surrounding rock loses its bearing capacity at a distance of 0.4 m from the surface of the rib.

(2) Borehole on left rib: obvious rock slip and dislocation are seen at 2.5 m from the orifice. A large scale fracture zone appears at a depth of 1.3 m. Furthermore, the mudstone is seriously damaged. The surrounding rock close to the rib (with a distance of 0.6 m) is extremely broken, without bearing capacity.

(3) Borehole on the floor: circumferential cracks are well developed at a depth of 3.8 m. Within 3.0 m from the orifice, the surrounding rock loses its integrity, with no bearing capacity.

The observations of the drilling holes show that within 3.0 m from the orifice, the integrity of the rock of the floor is extremely poor due to impact damage. In contrast, the maximum depth of the fracture zone on the right and left ribs is 3.6 m and 2.5 m, respectively. This indicates that compared to ribs, the floor is more vulnerable to dynamic load impact. Therefore, new control methods must be proposed to prevent the occurrence of heave.

3. Failure Criterion of Surrounding Rock Caused by Rock Burst

3.1. Stress Criterion

3.1.1. Static Load Analysis

By considering the correction coefficient [27], the shape of the roadway is equivalently simplified to a circular roadway. Meanwhile, the surrounding rock is reckoned as an elastic, homogeneous, and isotropic medium. Thereafter, the static load distribution of the rock around the roadway is denoted in Figure 3.

Usually, the equivalent radius of an irregular roadway is calculated with Equation (1):

$$R^{*}=\xi \sqrt{\frac{S}{\pi }}$$

(1)

where ξ refers to the correction coefficient. Given the shape of the No. −830 roadway, it is determined as 1.1 herein. S means the sectional area of the roadway, with a unit of m².

According to the principle of elasticity and the superposition of the static load stress fields, the static stress at any point of the surrounding rock of the equivalent circular roadway (in polar coordinates) is obtained with Equation (2) [28]:

$${\sigma }_{\mathrm{s}}=\{\begin{array}{l}{\sigma }_r=\frac{p}{2}\left[(1+\lambda )\left(1-\frac{R{*}^2}{r^2}\right)+(\lambda -1)\left(1-4\frac{R{*}^2}{r^2}+3\frac{R{*}^4}{r^4}\right)\cos 2\theta \right]+p_x\frac{R{*}^2}{r^2}\\ {\sigma }_{\theta }=\frac{p}{2}\left[(1+\lambda )\left(1+\frac{R{*}^2}{r^2}\right)-(\lambda -1)\left(1+3\frac{R{*}^4}{r^4}\right)\cos 2\theta \right]-p_x\frac{R{*}^2}{r^2}\end{array}$$

(2)

3.1.2. Dynamic Load Analysis

The coal and rock mass around the roadway are assumed to be elastic and isotropic and the stress wave propagates along the radial direction of the roadway. Therefore, the dynamic load generated by the stress wave propagation is calculated as Equation (3) [29]:

$${\sigma }_{\mathrm{d}}=\{\begin{array}{l}{\sigma }_{\mathrm{dP}}=\rho {\nu }_{\mathrm{P}}{\left({\nu }_{\mathrm{pp}}\right)}_{\mathrm{P}}{e}^{-\eta d}\\ {\sigma }_{\mathrm{dS}}=\rho {\nu }_{\mathrm{S}}{\left({\nu }_{\mathrm{pp}}\right)}_{\mathrm{S}}{e}^{-\eta d}\end{array}$$

(3)

where, ${\sigma }_{\mathrm{dP}}$ and ${\sigma }_{\mathrm{dS}}$ refer to the dynamic load caused by the propagation of P- and S-waves, while ${\nu }_{\mathrm{P}}$ and ${\nu }_{\mathrm{S}}$ are the propagation velocities. ${\left({\nu }_{\mathrm{pp}}\right)}_{\mathrm{P}}$ and ${\left({\nu }_{\mathrm{pp}}\right)}_{\mathrm{S}}$ represent the peak vibration velocity of particles caused by P-wave and S-wave propagation. η means the attenuation coefficient and d is the propagation distance.

The rock mass close to the No. 7102 workface experiences high static load stress. When the dynamic load induced by the mining activity is superimposed on this zone and reaches the critical stress, floor burst will occur. The schematic diagram of the superposition of dynamic and static loads [18] is shown in Figure 4. The failure criterion of the floor under the load superposition is written as Equation (4) [17]:

$$\sqrt{{\sigma }_r^2+{\sigma }_{\theta }^2}+{\sigma }_{\mathrm{d}}-{\sigma }_{\mathrm{b}min}\ge 0$$

(4)

where, ${\sigma }_{\mathrm{b}min}$ is the critical load (minimum load) in case of dynamic disaster.

From the point of view of stress, the main factors of impact damage on the roadway are the lateral pressure coefficient, distance of the earthquake source from the center of the roadway, dynamic load strength, depth of the roadway, and bearing strength of the surrounding rock. When the depth of the roadway and lateral pressure coefficient are large, the vibration intensity is high, and the source is close to the roadway, it is easier to induce impact damage in the rock surrounding the roadway.

3.2. Energy Criterion

According to the energy criterion, rock burst is caused by the accumulation of the large amounts of elastic energy in the surrounding rock, and the accumulated elastic energy is determined as Equation (5) [30]:

$$E_0=\frac{{\sigma }_1^2+{\sigma }_2^2+{\sigma }_3^2-2\mu \left({\sigma }_1{\sigma }_2+{\sigma }_1{\sigma }_3+{\sigma }_2{\sigma }_3\right)}{2E}$$

(5)

where μ represents Poisson’s ratio while E is the elastic modulus.

The rock mass is in a three-dimensional stress state before the coal is excavated and the accumulated elastic energy increases with the earth pressure. Under the influence of dynamic load caused by mining activity, the stress state of the surrounding rock changes and large amounts of accumulated energy is released and spread around in the form of stress waves. The released elastic energy is written as Equation (6):

$$E_{\mathrm{s}}=E_0-E_1$$

(6)

where E₁ is the residual elastic energy in rock mass and E_s represent the released energy.

The elastic energy is usually attenuated in exponential form while propagating; when the stress waves reach the surface, the energy left on the unit area is calculated as Equation (7):

$$E_2=E_{\mathrm{s}}{R}^{-\eta }=\left(E_0-E_1\right)R^{-\eta }$$

(7)

Where R is the distance between the epicenter and the surface of the floor/rib.

The energy absorbed by the original support is denoted as E₃ and the residual energy E₄ acting on the rock mass is:

$$E_4=E_2-E_3=\left(E_0-E_1\right)R^{-\eta }-E_3$$

(8)

The minimum energy consumed when floor burst occurs [17] is

$$E_{min}=\frac{{\sigma }_{\mathrm{b}min}^2}{2E}$$

(9)

Therefore, floor burst occurs when the residual energy is larger than the minimum energy consumed, which is expressed as Equation (10):

$$E_4\ge E_{min}\Rightarrow \left(E_0-E_1\right)R^{-\eta }-E_3\ge \frac{{\sigma }_{\mathrm{b}min}^2}{2E}$$

(10)

From the point of view of energy, the main factors affecting the degree of impact damage on the roadway are the energy of the source, the distance from the source, the physical properties of the surrounding rock, the energy absorption capacity of the support system, and other factors.

4. Failure Mechanism of the Surrounding Rock Caused by Rock Burst

4.1. Numerical Model

The Trigon model is a modeling method built in to the UDEC software. It cuts the traditional Tyson polygonal block into several triangular blocks to build the model, which better reveals the failure process by simulating the fracture, movement, and large deformation of the roadway. In the Trigon model, the coal-rock masses are divided into triangular blocks and connected together by contact surfaces between the blocks. Being elastic, each triangular block is divided into triangular finite difference regions where failure is unlikely to happen. The failure of coal-rock masses can only occur along the joint plane with shear or tensile failure. Figure 5 is the mechanical calculation criterion between the blocks in the Trigon model [31]. The Trigon model is closer to the real fractured coal-rock masses. Because the randomly generated triangular blocks and the contact surfaces can be better connected, the UDEC model working on triangular blocks can overcome the limitations of adopting parallel blocks, and thus can be used to study the instability and failure mechanisms of fractured coal-rock masses under different stress conditions in coal mining.

The variation relationship between stress and strain is represented by the value of $k_n$ [32]:

$$\Delta {\sigma }_n=-k_n\Delta u_n$$

(11)

where, $\Delta {\sigma }_n$ is the effective normal stress increment and $\Delta u_n$ is normal displacement increment. There is a limited tensile strength ${\tau }_{\mathrm{s}}^{max}$ for the contact. If the tensile strength is exceeded, then $\Delta {\sigma }_n$ = 0.

In the shear direction, the response is governed by a constant shear stiffness. The shear stress ${\tau }_{\mathrm{s}}$ is determined by a combination of contact micro properties, cohesion, and friction. Thus, if:

$$\left|{\tau }_{\mathrm{s}}\right|\le c+{\sigma }_n\tan \phi ={\tau }_{\mathrm{s}}^{max}$$

(12)

then

$${\tau }_{\mathrm{s}}=-k_{\mathrm{s}}\Delta u_{\mathrm{s}}^e$$

(13)

or else, if

$$\left|{\tau }_{\mathrm{s}}\right|\ge {\tau }_{\mathrm{s}}^{max}$$

(14)

then

$$\left|{\tau }_{\mathrm{s}}\right|=sign\left(\Delta u_s\right){\tau }_{\mathrm{s}max}$$

(15)

where c is the cohesion, φ is the friction angle, $\Delta u_{\mathrm{s}}^e$ is the elastic component of the incremental shear displacement, and $\Delta u_s$ is the total incremental shear displacement.

Based on the actual geological engineering conditions on site, the authors decided to build a numerical model of length × width = 120 m × 100 m with the discrete element software UDEC and analyze the failure and deformation features of the No. −830 main roadway under the dynamic load impact caused by the No. 7102 workface with the original support design. The designed roadway excavation width is 5.0 m, the vertical wall height is 1.8 m, and the vault height is 2.5 m. The calculation model is shown in Figure 6.

In order to improve the calculation efficiency, the UDEC Trigon logic model can be adapted to simulate real fractured coal-rock mass and thus is used to randomly divide the triangle blocks (with an average size of 0.2 m) around the roadway, so as to more accurately capture the fracture expansion and failure characteristics of the surrounding rock around the roadway, while the rock mass far away from the roadway is treated as conventional rectangular blocks with different aspect ratios. Specifically, the blocks in the model form linear elastic structures, whose structural surfaces meet the Coulomb slip criterion with residual strength, through the adoption of the built-in cable unit of UDEC as the anchor rod and the anchor cable, and the structural unit simulation of the ladder beam and shotcrete.

Table 1. Properties of support elements used in the model.

Support Elements	Contact Properties	Value
Cable	Cohesive strength of shear coupling spring (MPa)	1.2
	Stiffness of shear coupling spring (GPa)	8.4
	Frictional resistance of the shear coupling spring (°)	8.3
	Cohesive strength of normal coupling spring (MPa)	215
	Tensile yield strength (kN)	460
	Stiffness of normal coupling spring (GPa)	21
Bolt	Cross-sectional area(m2)	3.8 × 10−4
	Elastic modulus (GPa)	200
	Tensile yield strength (kN)	198
Structure	Elastic modulus (GPa)	200
	Tensile yield strength (MPa)	500
	Compressive yield strength (MPa)	500
	Interface normal stiffness (GPa/m)	10
	Interface shear stiffness (GPa/m)	10

shows the setting of mechanical parameters of the cables and other supporting components in the model. The boundaries on the left and the right ribs and the bottom of the model were fixed with the replacement method and a uniform load of 19.87 MPa equivalent to the weight of the overburden was applied on the top of the model.

According to the elastic wave theory [33], any complex stress wave can be obtained by the Fourier transform of several simple harmonics. Lu [34] pointed out that when the rock burst occurred on site, the vibration duration recorded by the microseismic monitoring system was only tens of milliseconds, and there was generally no effect of multiple rounds of rock burst. If the shock vibration source wave is simplified to a simple harmonic (a semisine wave) [26], the main frequency of the mine earthquake is about 10–20 Hz with a period of 0.05–0.1 s [35,36]. The relationship between the peak particle velocity (PPV) and the depth of the roadway [37] is shown in

Table 2. The relationship between the roadway and the PPV.

PPV/m/s	Depth of the Roadway (H)/m
PPV ≤ 0.05	H ≤ 300
0.05 < PPV ≤ 0.20	300 < H ≤ 500
0.20 ≤ PPV ≤ 0.40	500 < H ≤ 700
PPV > 0.40	H > 700

.

According to the microseismic monitor, the epicenter of this earthquake was on the mining face of the No. 7 coal seam of the roadway floor, which is at a horizontal distance of 30 m from the roadway. Based on the previous analysis, the impact dynamic load is simplified to a simple harmonic wave and the dynamic load causing the impact failure of the rock mass lasted for approximately one period (0.05 s), with a frequency of 20 Hz and an action time of 0.2 s, which was simulated dynamically by the built-in dynamic module in the UDEC software.

The waveform of the dynamic load is shown in Figure 7, and the dynamic strength is calculated as Equation (16):

$${\sigma }_{\mathrm{d}}=\rho C\upsilon$$

(16)

where ρ is the density of the propagation medium, C is the speed of wave propagation in the medium, and v is the peak particle velocity.

The model boundary was set to a viscous boundary to simulate the boundary in the infinite distance, and the stress wave was applied within 3 m of the model boundary on the right rib of the roadway.

The simulation was carried out in four steps:

Step 1: Imposing initial ground stress conditions in the global model, calculating the balance, and saving it as a numerical model of the original rock stress balance;

Step 2: Adopting the original rock stress balance model saved in the first step, deleting the internal block elements of the roadway contour and calculating the balance of the bolt and cable support, simulating the roadway driving process, and saving the model calculation results;

Step 3: Adopting the calculation results of the model saved in the second step, deleting some blocks in the No. 7 coal seam step by step, calculating the balance to simulate the impact on the No. 7102 working face, and saving the calculation results;

Step 4: Adopting the model calculation results saved in the third step, changing the boundary conditions of the model and applying a stress wave, and simulating the impact response characteristics of the roadway to the dynamic load.

4.2. Correction of Model Parameters

The intact rock parameters obtained by laboratory compression tests and Brazilian tests on standard specimens are listed in

Table 3. Intact rock properties and calculated rock mass properties.

Rock Strata	Intact Rock		RQD	Rock Mass
	Er/GPa	σc/MPa		Em/GPa	σcm/MPa	σtm/MPa
Sandstone	22.1	56.3	93	14.6	45.1	4.51
Mudstone	9.2	25.8	78	3.2	12.3	1.23
Sandy mudstone	14.1	36.4	85	6.6	21.4	2.14
No. 7 coal seam	3.2	12.2	70	0.8	4.62	0.46

.

Based on the physical and mechanical measurements of the surrounding rock, the RQD value obtained from drilling and peeking is reduced, and the following equation is used to modify the physical and mechanical parameters of the coal and rock mass used in the model [38,39].

$$\frac{E_m}{E_r}={10}^{0.0186\mathit{RQD}-1.91}$$

(17)

$$\frac{{\sigma }_{cm}}{{\sigma }_{\mathrm{c}}}={\left(\frac{E_m}{E_r}\right)}^q$$

(18)

where E_m and E_r represent the elastic modulus of the rock mass and the rock, respectively, σ_cm and σ_c show their uniaxial compressive strengths, respectively, q is the empirical parameter, and the calculation model is determined to be 0.7.

The tensile strength of the coal and rock mass (σ_tm) was estimated to be one-tenth of the compressive strength [40]. The properties of the rock mass were calculated and are listed in Table 3.

The normal stiffness and the shear stiffness of the structural plane are determined by the following equation:

$$k_n=n\left[\frac{K+4/3G}{\Delta {\mathrm{z}}_{\min }}\right],\left(1\le n\le 10\right)$$

(19)

$$k_{\mathrm{s}}=0.2k_n$$

(20)

where, k_n and k_s demonstrate the normal stiffness and shear stiffness of the structural plane, respectively, Δz_min means the minimum length of the contact area along the normal direction, K is the bulk modulus of the mass (K = E/3(1 − 2μ)), G displays the shear modulus of the mass (G = E/2(1 + μ)), and μ indicates Poisson’s ratio.

However, these parameters cannot be directly applied to the numerical model because numerical simulation predicts the natural deformation characteristics of the engineered rock mass. The difference of the physical and mechanical parameters between the engineered rock mass and the intact rock measured in the laboratory cannot be ignored. If the parameters measured in the laboratory are converted and directly substituted into the calculation model, the results are bound to be unreliable.

For the sake of reducing the influence of the blocks size effect and obtaining the accurate mechanical parameters of different rock masses in the numerical calculation model, the authors used the trigon logic model to establish a correction model so as to conduct UCS and BTS experiments on rock masses with dimensions of 2 m wide × 4 m high [31] (Figure 8a) and 2 m diameter [41] (Figure 8b), in line with the proportions of the block size. Since the models are composed of four kinds of lithology, four contact types must be calibrated. The determined parameters of coal and rock mass, as well as contact surface in the numerical model, are displayed in

Table 4. Physical and mechanical parameters of blocks and contacts in model.

Rock Strata	Rock Mass Properties			Contact Surface Parameters
	Dens/kg/m3	Bulk/GPa	Shear/GPa	Jkn/GPa·m−1	Jks/ GPa·m−1	Jcoh/MPa	Jfric/°	Jtens/MPa
Sandstone	2 600	10.15	5.80	23.6	4.72	2.7	32/24	2.1
Mudstone	2 100	2.07	1.30	14.2	2.84	1.2	28/20	0.9
Sandy mudstone	2 350	4.79	2.60	18.5	3.70	1.6	23/18	0.6
No. 7 coal seam	1 300	0.53	0.32	6.1	1.22	0.9	21/15	0.2

. The corrected results of numerical calculation model are shown in Figure 8c,d and

Table 5. Comparison of theoretical and simulated values of rock mass parameters.

Rock Strata	E/GPa		Error/ 100%	UCS/MPa		Error/ 100%	BTS/MPa		Error/ 100%
	Target	Calibrated		Target	Calibrated		Target	Calibrated
Sandstone	14.60	14.49	−0.75	45.1	45.8	1.55	4.51	4.64	2.80
Mudstone	3.20	3.14	−1.88	12.3	12.4	0.08	1.23	1.18	−4.07
Sandy mudstone	6.60	6.72	1.81	21.4	20.8	−2.80	2.14	2.10	−1.86
No. 7 coal seam	0.80	0.75	−6.25	4.62	4.60	−0.04	0.46	0.49	6.52

.

It can be seen from Table 5 that the error between the uniaxial compressive strength and elastic modulus obtained by numerical simulation and the data obtained by laboratory test is within 10% [42,43]. Therefore, the mechanical parameters of coal and rock mass determined in Table 5 are reasonable.

4.3. Simulation Result

After running the numerical simulation, the stress, surface displacement, and floor displacement vectors of the surrounding rock of the No. −830 main roadway under the original support scheme were analyzed and the results are as follows.

(1) Stress evolution.

As shown in Figure 9, (i) during the roadway excavation, due to the superposition of dynamic and static loads, the stress at the surface of the rock surrounding the roadway was significantly released, and the stress release area of the two ribs and the floor was comparatively larger, while at the interface between mudstone and sandstone, stress showed differential release characteristics. (ii) Under the influence of mining, the surface stress on the roadway was further released. Under the influence of the lateral supporting pressure of the working face, the stress concentration phenomenon on the roof and floor was further increased and the concentrated stress reached up to 32.15MPa. Under the effect of the dynamic load disturbance caused by the fracture of the rock formation, the concentrated stress of the roof and floor of the roadway transferred to the deep part, and the increase was relatively apparent. (iii) Compressive stress concentration occurred again at 4.2 m on the roof of the roadway and 6.5 m on the floor. The maximum horizontal compressive stress was 39.64 MPa. According to the stress evolution characteristics, the impact damage of dynamic load on the roadway floor was relatively obvious under a high stress environment. The higher the level of stress concentration, the greater the degree of floor damage.

(2) Fracture propagation.

As shown in Figure 10: (i) during the roadway excavation, the shallow part of the surrounding rock in the in-situ stress state was relatively obviously damaged by tension. As the stress shifted in depth and tended to a balanced state, the development of cracks on the surface of the surrounding rock temporarily slowed down. The rock surrounding the roadway maintained a relatively stable state under the action of the bolt and cable active support system. (ii) During the mining of the working face, the rock mass in the high static stress environment further deteriorated the stress environment of the rock surrounding the roadway. The roadway floor failure and deformation were relatively obvious when the rock mass at the bottom corner was completely within the crack expansion range. As the degree of development of the surrounding rock fissures further increased, the relatively stable state of the rock masses surrounding the roadway was broken again, and the roadway floor began to bulge and slowly rise. (iii) When the roadway was impacted by the dynamic load, the mudstone tension cracks in the roadway floor expanded further and the scope of the slip shear cracks expanded. The cracks developed and caused damage to the surrounding rock structure to increase within the bolt anchorage range. The anchoring system basically lost its active bearing capacity and the broken rock mass continued to shift to the empty area of the roadway. Therefore, dynamic load disturbance is the inducement for the frequent occurrence of floor heave after routine repair of the roadway.

(3) Displacement distribution.

It can be seen from Figure 11 that: (i) when the floor is unsupported, the integrity damage of the surrounding rock structure by the driving action is weak, and the deformation of the rock surrounding the roadway is not obvious compared with the influence of mining and the disturbance of the dynamic load. (ii) During mining on the workface, the accumulated displacement of the roadway floor and the two ribs increased by 340 mm and 583 mm, respectively, compared with the deformation of the surrounding rock during the excavation. (iii) Under the action of dynamic load disturbance, the cumulative maximum floor heave volume of the roadway is 1248 mm, which is 1.99 times the cumulative floor heave volume of the roadway during the mining period. Therefore, floor heave of the roadway is an obvious feature after the roadway is impacted by the dynamic load, which is consistent with the actual deformation characteristics of the site.

Figure 12 is a vector diagram of the overall displacement of the rock surrounding the roadway. It can be seen that: (i) under the condition of no support on the floor, the disturbance of the road excavation causes the yield deformation of the shallow rock mass in the two ribs of the roadway and the bottom corner area, and transfers to the roadway surface. (ii) During the mining of the working face, as the disturbance effect increases, the broken rock mass begins to flow to the empty area of the roadway, and the moving speed increases plastically. (iii) When the roadway is impacted by the dynamic load, the stress on the surrounding rock under the limit equilibrium state is increased, and the original equilibrium state is broken. Under the influence of the continuous increase in the subsidence of the two ribs of the roadway and the increasing damage of the bottom corner, the rock mass in the deep part of the floor quickly shifts to the shallow part of the floor, resulting in serious floor heave of the roadway.

The conventional repair of “floor-brushing and rib-brushing” on the mudstone floor of the No. −830 main roadway cannot prevent floor deformation because of the high in situ stress, as well as the existence of dynamic loads. Instead, frequent repair of the floor and rib will speed up the failure of the surrounding rock, which will further result in serious floor burst.

4.4. Failure Mechanism of the Roadway

Based on the results of field investigation, theoretical analysis, and numerical simulation, the deformation and failure mechanism of the No. −830 main roadway is summarized in Figure 13.

(1) The influence of superposition of the dynamic and static loads.

Considering that the average buried depth of the No. −830 main roadway is 870 m, it is subjected to high in situ stress. In addition, the strong dynamic load caused by the advancing of the No. 7102 workface makes large deformation of the surrounding rock more possible. Therefore, the superposition of the static and dynamic loads is the major trigger of floor burst.

(2) The difference in the strength of surrounding rocks.

The roof and upper-side rib of the roadway are made of sandstone, which is characterized by comparatively higher strength, while the majority of the floor and lower-side rib are mudstone. Given the lower strength and cohesion of mudstone, a large range of fracture and plastic zones appear in the lower-side rib and floor after the application of the dynamic load. In addition, the failure of the initial support and the obvious movement of the rib (due to the sliding effect along the bedding plane) weaken the supporting effect of the two ribs, which further reduces the bearing capacity of the floor, accelerating the occurrence of large deformation of the surrounding rock and floor burst.

(3) Insufficient supporting strength.

The two ribs and roof of the roadway are initially supported by shotcrete. Thereafter, bolt-mesh supporting is adopted to further stabilize the roadway. In contrast, no supporting measure is applied on the floor, which makes the floor the weakest plane of the roadway. The results of the numerical simulation indicate that when the bolt is anchored in the fracture zone, it results in insufficient anchoring force and, as a result, larger deformation occurs. This in turn speeds up the subsidence of the two ribs. Moreover, the superposition of the static and dynamic loads leads to significant plastic deformation and even sliding shear failure of the floor. The accumulated energy is more likely to be released through floor failure (floor burst).

(4) Apparent hydrolysis weakening.

Mudstone tends to swell and soften with water. According to the analysis of laboratory tests, the strength of mudstone decreases by 51.75% after being soaked for 48 h. Because the drainage ditch and other waterproof measures in the No. −830 roadway are not well prepared, fissure water is observed in mudstone. Long-term immersion in water has aggravated the softening of the mudstone. Consequently, the mechanical properties of mudstone are severely degraded, making the floor more vulnerable to in situ stress.

5. Control Technology of Rock Burst in Roadway

5.1. Methodology

Given the stress distribution, failure characteristics, internal damage of surrounding rock, and the numerical simulation results under the original support, a combined support method which includes anchor cables and arched girders was used to reinforce the floor in No. −830 roadway. The specific processes are introduced below:

5.1.1. Full Section Anchor Cables

According to the results obtained from borehole peep and numerical analysis, it was known that the surrounding rock around the roadway was highly damaged and fracture was seriously propagated. Besides, due to the high dynamic load, some of the anchor bolts were losing their supporting capacity. Therefore, pre-stressed anchor cables were set in the full section roadway to restrict the development of the plastic zone. In addition, the adoption of anchor cables can also increase the length of anchorage sections and improve the stability of the roadway, even under the superposition of static and dynamic loads. Furthermore, the anchor cables installed on the floor can cut off the plastic slip line of the floor and prevent movement of the Langken zone.

5.1.2. Usage of Arched Girders

After the occurrence of rock burst, large amounts of rocks were pushed to the roadway from the rib corner. When the arched girder was applied, it formed a combined bearing structure with damaged rock (in the shallow area) and intact rock (in the deep zone). Specifically, the arched girder can enhance the flexural capacity of the floor and stop the broken rock from moving to the roadway. Meanwhile, the damaged rocks in the shallow area act as a buffer zone that weakens stress waves and energy during propagation. This eliminates the damage to the floor caused by dynamic load. The main function of the intact rock in the deep zone is that it proves a strong base for the arched girder. In other words, the arched girder is fixed to the intact rock by the anchor cables. This combined bearing structure can effectively restrict the deformation (due to the superposition of static and dynamic loads) of the floor.

5.2. Control Scheme

The scheme of the combined support method is shown in Figure 14.

Two kinds of anchor cables (long cable and short cable) were used. The specifications of the short cable were 18.9 mm diameter and 4300 mm length. The size of the pallet that goes with the short cable was 250 mm × 250 mm × 14 mm. The diameter of the long cable was the same as short one, but the length was 7300 mm, which is longer than the short cable. The pallet for the long cable had a size of 300 mm × 300 mm × 14 mm. As can be seen from the figure, the rod and row spacing of the long cable were 2000 mm and 1000 mm. In contrast, the rod and row spacing of the short cable at the roof were both 1000 mm, while those installed for the rib and floor stabilization were 800 mm and 1000 mm. The types of anchoring agents used for cables were Z2370 and CK2350 and the pre-stress was no less than 150 kN.

The field test and numerical simulation indicate that the maximum amount of floor heave is 1.25 m. The fracture zone under the floor has a depth of 3.8 m, while the depth of the fracture zone on the two ribs is 0.5 m. Therefore, after initial floor brushing (with a depth of 1.25 m), another 1.75 m of fracture zone needed to be removed and the residual 0.8 m of damaged rock mass was left as the buffer zone. The cement, C30, was used to stabilize the arched girder. Considering the potential swelling ability of the mudstone, lime was adopted before the construction of the arched girder to stop fissure water from having access to the floor.

5.3. Engineering Application

After the application of the combined support method in No. −830 roadway, several monitoring spots were arranged to record the displacement of the surrounding rock. Figure 15a shows the monitoring results of the deformation up to 90 days.

From the results, it can be seen that the roof and ribs experienced large deformation within 20 days. After the initial significant movement, the roof and ribs deformed gradually (from the twentieth day to fiftieth day) due to the restriction posed by anchor cables. In contrast, the floor showed fast heave from the thirtieth day, and after 60-days of rising, the floor heave stabilized at 67.9 mm. This is mainly attributed to dynamic load caused by the advancing of the No. 7102 workface.

The stability of the roadway was well maintained after the application of the combined support method. The maximum displacement of two ribs was 82.8 mm, which is 18% of the deformation with the original support. Moreover, the maximum subsidence of the roof was 48.5 mm, which is a decrease of 95% compared to the initial subsidence. The actual scene of the roadway after the application of the combined support method is shown in Figure 15b, from which it is concluded that the usage of anchor cables and arched girders worked well to restrict deformation in the roadway. This provides meaningful guidance when it comes to the support of roadways with similar geological conditions.

6. Conclusions

This manuscript provides specific support reinforcement techniques for the control of deformation (caused by rock burst) of the weak floor, and the following conclusions were obtained:

(1)

The superposition of high static and strong dynamic loads, prominent differences in the strength of the surrounding rock, poor mechanical qualities of the floor, and insufficient support strength of the ribs and the floor are the main reasons for the rock burst of the floor.

(2)

In the environment dominated by high static load in the near field, fractures in the surrounding rock developed and strength was reduced; under the influence of the lateral stress of mining, the deformation of the surrounding rock increased; with the action of dynamic load disturbance in the far field, the impact damage on the soft rock of the roadway floor was serious.

(3)

This paper proposes a combined support method which include anchor cables and inverted arches. The on-site monitoring results indicate that the maximum floor heave is about 67.9 mm with the combined support technique, which is 95% lower than that under the original support conditions. The roadway floor burst was under well control.