Research on Floor Heave Mechanisms and Control Technology for Deep Dynamic Pressure Roadways

Abstract

In order to study the influencing factors of floor deformation and floor heave mechanisms of deep mining roadways, this paper takes the deep dynamic pressure mining roadway of a mine as the engineering background and adopts a research method combining theoretical analyses, numerical simulations and field observations to study the influence of various factors on floor deformation and floor heave mechanisms. It is determined that the influencing factors on floor heave are a large buried depth, a long duration of dynamic pressure, unique characteristics of the surrounding rock and an insufficient support strength. A bearing mechanics model of the roadway floor beam is established, and it is determined that the displacement of the roadway floor is negatively correlated with the elastic modulus and floor thickness and positively correlated with the buried depth of roadway, the roadway width and the width of fracture zone. A numerical simulation method is used to study the influence of the original geological conditions, strengthening the elastic modulus of floor, strengthening the strength of the side wall rock and increasing the thickness of the floor rock on the displacement of the roadway floor. It is determined that increasing the thickness of floor rock controls the floor heave the most, followed by strengthening the elastic modulus of the floor rock and then strengthening the strength of the side walls. The results of the numerical simulation agree well with those of the theoretical analysis. After the control method of “bottom lifting + bottom angle bolt + floor bolt ” is adopted on site to treat the floor heave, the floor heave volume of the roadway is small during the service period of the 303 working face return air roadway, which meets the application requirements of the roadway. Meanwhile, the theoretical analysis and numerical simulation results are indirectly verified.

Roadways are the passages that transport underground coal resources from mining, and the surrounding rock stability directly affects the safe and efficient production of the mine [1]. According to the relevant statistics, floor heave failures account for a large proportion of the deformation and failure of the surrounding rock of roadways [2,3]. With the increase in mining depth for coal resources, the influence of a “four high and one disturbance” mining environment on roadways is introduced [4] and the deformation of the surrounding rock is rapidly increased. In particular, the phenomenon of floor heave in roadways is more prominent, which causes an increase in repair times and secondary maintenance costs of roadways and has become a key hidden danger, threatening the stability of the surrounding rock in deep roadways. At the same time, the traditional concept of attaching importance to roof and side surrounding rock support and disregarding floor support has been difficult to adapt in the control of the surrounding rock in deep roadways, especially under the action of dynamic pressure in working face mining. How to control the stability of roadway floors under deep dynamic pressure has become the key problem in deep resource mining.

In view of the floor heave mechanisms of deep roadways, researchers have divided floor heave into several types: extrusion flow type, flexural fold type, shear disjoint type and water expansion type [5,6,7]. Researchers have analyzed the causes of all kinds of floor heave [8] and proposed corresponding control measures, for example, the construction of a bottom corner anchor bolt [9], over-excavation anchor grouting backfill [10], grooving pressure relief [11], reinforcement of roadway sides [12,13], reinforcement of floors with a full length anchored hydraulic expansion anchor bolt [14], anchor mesh shotcrete and delayed grouting [15], stress transfer [16], full face grouting and side angle anchor bolt [17], concrete inverted arches and prestressed anchor cables. As a special roadway, dynamic pressure roadways have experienced excavation disturbance, mining disturbance of the adjacent working face and mining disturbance of its own working face, and the surrounding rock deformation and failure process is different from that of conventional roadways [18,19]. Centering on the failure characteristics and stability control measures of the surrounding rock of dynamic pressure roadways, researchers have proposed the full anchor cable reinforcement technologies of short anchor cable support and long anchor cable reinforcement [20], high resistance pressure yielding support technology [21], anchor net cable and anchor injection combined support [22] and non-uniform anchor injection reinforcement [23,24].

The above studies have revealed the mechanism of floor heave from different angles and proposed corresponding floor heave control measures and surrounding rock stability control methods for dynamic pressure roadways. However, most studies were based on single factors such as the deep mining stress environment, the special geological conditions and the complex working conditions, and few have involved the study of floor heave under multi-factor coupling. Based on this, this paper takes the dynamic pressure roadway under the deep mining environment of a mine as the research object; comprehensively adopts a research method combining theoretical analysis, numerical simulation and field observations; discusses the influencing factors of deep dynamic pressure roadway floor heave; reveals the mechanism of floor heave; formulates the corresponding control measures for floor heaving; and provides references for similar roadway floor stability control under the multi-factor coupling effect.

1. Project Overview

The average mining depth of a 301 working face in a mine is 750 m, which is classified as deep mining. The coal seam has an average thickness of 4.7 m and an average dip angle of 1°. The working face adopts a three-lane system layout of belt haulage drift, auxiliary haulage drift and return air drift. The roadway layout of the working face is shown in Figure 1.

As can be seen from Figure 1, the auxiliary transportation roadway of a 301 working face has been successively used as the auxiliary transportation roadway of a 301 working face and the return air roadway of a 303 working face, and has experienced the influence of driving disturbance, resulting in serious deformation of the surrounding rock. It is classified as a deep dynamic pressure roadway.

The section size of the auxiliary transportation roadway in a 301 working face is 4.6 m × 3.8 m, and the combined support of bolting with wire mesh and cable is adopted. The support parameters are as follows: (1) Φ22 × 2800 mm left-hand thread steel bolt without a longitudinal bar is used in the roof, with a row spacing of 650 mm × 800 mm. One MSK23/35 and two MSZ23/60 resin anchoring agents are used for a single anchor rod. Φ21.8 × 10,300 mm anticorrosive anchor cables are used, with a row spacing of 1300 mm × 800 mm. One MSK28/50 and three MSZ28/50 resin anchoring agents are used for a single anchor cable, and the anchor cable is arranged in the middle of two rows of anchor rods. The top plate is equipped with a 4200 mm × 5000 mm composite mesh. The anchor rod presses down on the reinforcement joist made of Φ16 round steel with a beam length of 4100 mm and a hole spacing of 650 mm. The anchor cable joist is made of a T140 steel strip with a length of 4100 m and a hole spacing of 1300 mm. (2) Φ22 × 2800 mm metal anchor rods are used on both sides, with a row spacing of 800 mm × 800 mm, and T100 steel strips with a length of 2600 mm are pressed down on the upper four rows of anchor rods. The side of the coal pillar is hung with Φ6.5 mm reinforcement mesh with a specification of 2000 mm × 1000 mm. The side of the mining face is equipped with 1200 mm × 3600 mm composite mesh. The roadway section support is shown in Figure 2.

According to the actual situation, the section shrinkage rate of the 301 auxiliary transportation roadway is large, and especially, the floor heave of the roadway is more serious during the mining of the 301 working face. After the stoping of the working face, the floor heave of the working face gradually increases along the auxiliary transportation roadway of the working face 301 from the stoping line position to the cutting hole position of the working face. The floor heave volume of roadway near the cut hole is close to 2.0 m; that is, the floor of auxiliary transportation roadway in the 301 working face is seriously affected by dynamic pressure disturbance and the deep stress environment. The central position of the auxiliary transportation roadway is shown in Figure 3.

According to the actual deformation of the auxiliary transportation roadway on the 301 working face, the section of the roadway after deformation is shown in Figure 4. The deformation of the roof, floor and surrounding rock of the roadway is significant, and the reduction in area is close to 50%. The deformation of the surrounding rock is mainly manifested in the side of the floor and coal pillar, especially on the floor. The entire floor moves up, and the two sides and the center of the floor break upwards in some areas. The roadway floor is seriously affected by the coupling of a deep stress environment and dynamic pressure disturbance.

2. Mechanism of Floor Heave in Roadway

2.1. Analysis of Influencing Factors

Research has shown that the factors affecting the stability of the surrounding rock of the roadway include the surrounding rock properties, the buried depth, the stress environment, support strength, etc., and the influence of each influencing factor on the roadway is not a single effect, but the result of comprehensive action with other factors [25]. Combined with the use and spand and ecific geological conditions during the service period of the auxiliary transportation roadway in the 301 working face, it is determined that the influencing factors on the floor heave are the large burial depth, the long period of dynamic pressure influence, the unique occurrence of surrounding rock and an insufficient support strength. The influencing factors are analyzed as follows:

(1) Large burial depth: The average buried depth of the auxiliary transportation roadway in the 301 working face is 750 m and the original rock stress reaches 18.75 MPa. At 1.2 times the lateral pressure coefficient [26], the horizontal stress of the surrounding rock of the roadway is about 22.5 MPa; the roadway is in a high stress environment as a whole. The plastic zone of coal with a low side strength increases, reducing the constraint effect on the floor. Furthermore, the actual span of the floor is increased and the self-stabilization ability of the floor is reduced.

(2) Long period of dynamic pressure influence: The auxiliary transportation roadway of the 301 working face has successively experienced driving disturbance, mining disturbance from the 301 working face and mining disturbance from 303 working face, which makes the surrounding rock of the roadway relatively broken. Especially after mining of the 301 working face, the lateral arc-shaped triangular block of the working face transfers the overlying rock load to the coal pillar, which then transfers the load to the floor, which makes the floor bear the load unevenly. In order to release the load of the overlying strata transferred by the coal pillar, the floor of the roadway is deformed because it is in the free surface.

(3) Surrounding rock occurrence is unique: The immediate roof of the 301 working face auxiliary transportation roadway is fine sandstone with a thickness of 2.8~3.7 m, which is basically siltstone with thickness of 11~27 m. The immediate bottom is mudstone with a thickness of 1.2~4.0 m, and the Prinell hardness is 3~4. The roadway floor is easy to weather and break, it swells when it meets water, and thus its strength decreases rapidly. Meanwhile, the mudstone of the roadway floor is thin and has obvious stratification, showing the deformation characteristics of a layered soft rock roadway [27].

(4) Insufficient support strength: The roof and the side of the auxiliary transportation roadway in the 301 working face are supported by a combined bolt–mesh–anchor support, while the floor, without support, is the weak link of the surrounding rock of the roadway. Under the comprehensive action of all the above factors, a large deformation will occur to release the load and energy gathered by the surrounding rock.

2.2. Base Plate Load Bearing Characteristics

According to the actual situation of the auxiliary transportation roadway in the 301 working face, it can be seen that the floor of the roadway is not supported and affected by the mining of the 301 working face. The wall rock of the coal pillar side of the roadway is relatively broken, and its plastic zone has a large development width. Therefore, the roadway floor can be approximated as a simply supported beam. The mechanical model of the unit length floor is established as shown in Figure 5.

In the Figure 5, L is the width of the roadway, d is the thickness of the floor, L₁ is the width of the fracture zone at the coal pillar side, L₂ is the width at the fracture zone at the solid coal side, q is the load intensity of the lower floor to the immediate floor, q₁ is the load concentration within the fracture zone range at the coal pillar side, q₂ is the load concentration within the fracture area range at the solid coal side and q₃ is the horizontal stress concentration.

It is assumed that the roadway floor is an isotropic elastic body, and the effects of each load on the floor are independent [28]; that is, the deflection of the roadway floor is equal to the sum of the deflections under the separate effects of each load. At the same time, it is assumed that the deformation of the roadway floor under horizontal stress is based on the deformation caused by other loads. In combination with the calculation principle of the superposition method, the load borne by the roadway floor is decomposed into the load concentration, q, of the lower floor to the immediate floor, the load concentration, q₁, within the fracture area of the coal pillar side, the load concentration, q₂, within the fracture area of the solid coal side and the horizontal stress concentration, q₃. The mechanical model under each load is shown in Figure 6.

The coordinate system shown in Figure 6 is established. The horizontal direction to the right is the x-axis, the vertical direction upward is the y-axis, and the upwards displacement of the floor beam is positive while downwards is negative. Combined with the theory of material mechanics, the bending moments of the beam at different positions under different loads are solved, and the displacement (heave) of the floor beam under different loads can be obtained according to the approximate differential equation of the flexural line. According to the bending moment calculation formula, the bending moments at position x from the origin under various loads can be obtained by Formulas (1)–(3):

$$\{\begin{array}{l}\mathrm{When} x\in \left[0, L_1+L+L_2\right]:\\ M_a\left(x\right)=\frac{q{x}^2-q\left(L_1+L+L_2\right)x}{2}\end{array}$$

(1)

$$\{\begin{array}{l}\mathrm{When} x\in \left[0, L_1\right]:\\ M_b\left(x\right)=\frac{q_1{L}_{1}x}{2}-\frac{q_1{L}_1{}^{2}x}{6\left(L_1+L+L_2\right)}-\frac{q_1{x}^2}{2}+\frac{q_1{x}^3}{6L_1}\\ \mathrm{When} x\in \left[L_1, L_1+L+L_2\right]: \\ M_b\left(x\right)=\frac{q_1{L}_1{}^2}{6}-\frac{q_1{L}_1{}^{2}x}{6\left(L_1+L+L_2\right)}\end{array}$$

(2)

$$\{\begin{array}{l}\mathrm{When} x\in \left[0, L_1+L\right]:\\ M_c\left(x\right)=\frac{q_2{L}_2{}^{2}x}{6\left(L_1+L+L_2\right)}\\ \mathrm{When} x\in \left[L_1+L, L_1+L+L_2\right]: \\ M_c\left(x\right)=\frac{q_2{L}_2{}^{2}x}{6\left(L_1+L+L_2\right)}-\frac{q_2{\left(x-L_1-L\right)}^3}{6L_2}\end{array}$$

(3)

The approximate differential equation of flexure is shown in Equation (4):

$$\frac{d^{2}y}{d{x}^2}=\frac{M\left(x\right)}{EI}\Rightarrow EI{d}^{2}y=M\left(x\right)d{x}^2$$

(4)

The formulas for calculating the displacement (d) of the floor rock beam under the action of each load can be obtained by substituting the Formulas (1)–(3) into Formula (4) as follows:

$$\{\begin{array}{ll}\mathrm{When}& x\in \left[0, L_1+L+L_2\right]:\\ y_a\left(x\right)& =\frac{q{x}^4}{24EI}-\frac{q\left(L_1+L+L_2\right)x^3}{12EI}\\ & +\frac{q{\left(L_1+L+L_2\right)}^{3}x}{24EI}\end{array}$$

(5)

$$\{\begin{array}{ll}\mathrm{When}& x\in \left[0, L_1\right]:\\ y_b\left(x\right)& =\frac{q_1{L}_1{x}^3}{12EI}-\frac{q_1{L}_1{}^2{x}^3}{36EI\left(L_1+L+L_2\right)}-\frac{q_1{x}^4}{24EI}\\ & +\frac{q_1{x}^5}{120EI{L}_1}+\frac{q_1{L}_1{}^{3}x}{24EI}-\frac{q_1{L}_1{}^2\left(L_1+L+L_2\right)x}{18EI}\\ & -\frac{q_1{L}_1{}^{4}x}{120EI\left(L_1+L+L_2\right)}\\ \mathrm{When}& x\in \left[L_1, L_1+L+L_2\right]: \\ y_b\left(x\right)& =\frac{q_1{L}_1{}^4}{120EI}+\frac{q_1{L}_1{}^2{x}^2}{12EI}-\frac{q_1{L}_1{}^2{x}^3}{36EI\left(L_1+L+L_2\right)}\\ & -\frac{q_1{L}_1{}^2\left(L_1+L+L_2\right)x}{18EI}-\frac{q_1{L}_1{}^{4}x}{120EI\left(L_1+L+L_2\right)}\end{array}$$

(6)

$$\{\begin{array}{ll}\mathrm{When}& x\in \left[0, L_1+L\right]:\\ y_c\left(x\right)& =\frac{q_2{L}_2{}^2{x}^3}{36EI\left(L_1+L+L_2\right)}+\frac{q_2{L}_2{}^{4}x}{120EI\left(L_1+L+L_2\right)}\\ & -\frac{q_2{L}_2{}^2\left(L_1+L+L_2\right)x}{36EI}\\ \mathrm{When}& x\in \left[L_1+L, L_1+L+L_2\right]: \\ y_c\left(x\right)& =\frac{q_2{L}_2{}^2{x}^3}{36EI\left(L_1+L+L_2\right)}-\frac{q_2{\left(x-L_1-L\right)}^5}{120EI{L}_2}\\ & +\frac{q_2{L}_2{}^{4}x}{120EI\left(L_1+L+L_2\right)}-\frac{q_2{L}_2{}^2\left(L_1+L+L_2\right)x}{36EI} \end{array}$$

(7)

where M(x) is the bending moment; E is the elastic model in GPa; I is the moment of inertia, where I = hd³/12; d is the thickness of the floor rock beam in m; h is the length of the rock beam, where h = 1.0 m; y_a, y_b and y_c are the displacements in m of (a), (b) and (c) in Figure 6, respectively, when each load is applied alone.

According to the principle of the superposition method, the formula for the displacement of the floor rock beam under the combined action of various loads can be obtained, and is shown in Formula (8).

$$\{\begin{array}{l}\mathrm{When} x\in \left[0, L_1\right]:\\ y\left(x\right)=\frac{1}{E{d}^3}\left(\begin{array}{l}\frac{q{x}^4}{2}-q\left(L_1+L+L_2\right)x^3+\frac{q{\left(L_1+L+L_2\right)}^{3}x}{2}+q_1{L}_1{x}^3-\frac{q_1{L}_1{}^2{x}^3}{3\left(L_1+L+L_2\right)}-\frac{q_1{x}^4}{2}+\frac{q_1{x}^5}{10L_1}+\frac{q_1{L}_1{}^{3}x}{2}\\ -\frac{q_1{L}_1{}^2\left(L_1+L+L_2\right)x}{1.5}-\frac{q_1{L}_1{}^{4}x}{10\left(L_1+L+L_2\right)}+\frac{q_2{L}_2{}^2{x}^3}{3\left(L_1+L+L_2\right)}+\frac{q_2{L}_2{}^{4}x}{10\left(L_1+L+L_2\right)}-\frac{q_2{L}_2{}^2\left(L_1+L+L_2\right)x}{3}\end{array}\right)\\ \mathrm{When} x\in \left[L_1, L_1+L\right]: \\ y\left(x\right)=\frac{1}{E{d}^3}\left(\begin{array}{l}\frac{q{x}^4}{2}-q\left(L_1+L+L_2\right)x^3+\frac{q{\left(L_1+L+L_2\right)}^{3}x}{2}+\frac{q_1{L}_1{}^4}{10}+q_1{L}_1{}^2{x}^2-\frac{q_1{L}_1{}^2{x}^3}{3\left(L_1+L+L_2\right)}\\ -\frac{q_1{L}_1{}^2\left(L_1+L+L_2\right)x}{1.5}-\frac{q_1{L}_1{}^{4}x}{10\left(L_1+L+L_2\right)}+\frac{q_2{L}_2{}^2{x}^3}{3\left(L_1+L+L_2\right)}+\frac{q_2{L}_2{}^{4}x}{10\left(L_1+L+L_2\right)}-\frac{q_2{L}_2{}^2\left(L_1+L+L_2\right)x}{3}\end{array}\right)\\ \mathrm{When} x\in \left[L_1+L, L_1+L+L_2\right]:\\ y\left(x\right)=\frac{1}{E{d}^3}\left(\begin{array}{l}\frac{q{x}^4}{2}-q\left(L_1+L+L_2\right)x^3+\frac{q{\left(L_1+L+L_2\right)}^{3}x}{2}+\frac{q_1{L}_1{}^4}{10}+q_1{L}_1{}^2{x}^2-\frac{q_1{L}_1{}^2{x}^3}{3\left(L_1+L+L_2\right)}-\frac{q_1{L}_1{}^{4}x}{10\left(L_1+L+L_2\right)}\\ -\frac{q_1{L}_1{}^2\left(L_1+L+L_2\right)x}{1.5}+\frac{q_2{L}_2{}^2{x}^3}{3\left(L_1+L+L_2\right)}-\frac{q_2{\left(x-L_1-L\right)}^5}{10L_2}+\frac{q_2{L}_2{}^{4}x}{10\left(L_1+L+L_2\right)}-\frac{q_2{L}_2{}^2\left(L_1+L+L_2\right)x}{3}\end{array}\right)\end{array}$$

(8)

It can be seen from the Formula (8) that the roadway floor displacement is affected by the elastic modulus, E, of the floor rock beam, the thickness, d, of the floor rock beam, the load concentration, q, of the lower floor to the immediate floor, the load concentration, q₁, within the fracture area of the coal pillar side, the load collection, q₂, within the fracture area of the solid coal side and the roadway width, L. The width of the fracture zone on the coal pillar side is L₁ and the width of the fracture zone on the solid coal side is L₂. The fracture zone is a part of the limit equilibrium zone of the roadway surrounding rock and is adjacent to the plastic zone of the roadway surrounding rock; therefore, the maximum load intensity in the fracture zone is equal to the original rock stress and the width of the fracture zone can be approximately solved, according to the reference [28]. Considering that the surrounding rock is in a static equilibrium state before the excavation of the roadway, the vertical stress at the immediate bottom of the roadway is equal to the load of the lower floor on the immediate bottom; that is, $q=q_0=\gamma H$. At this time, $q=q_1=q_2=\gamma H$, where γ is the overburden bulk density.

2.3. Influencing Characteristics of Each Factor

In order to determine the influence characteristics of each influencing factor on the roadway floor displacement, the single factor analysis method is used to study the change law of the floor displacement at different positions with each factor. The specific experimental scheme is shown in

Table 1. Experimental program and parameters.

Factor	Buried Depth	Thickness	Elasticity Modulus	Width of Roadway	Fracture Zone Width
	H (m)	d (m)	E (GPa)	L (m)	L1 (m)	L2 (m)
Parameter	300	1.5	15	4.0	1.0	0.5
	400	2.0	20	4.5	1.5	0.8
	500	2.5	25	5.0	2.0	1.0
	600	3.0	30	5.5	2.5	1.2
	700	3.5	35	6.0	3.0	1.5
	800	4.0	40	6.5	3.5	1.8

. In Table 1, different burial depths are used to replace each load, and the overburden bulk density is taken as 25 kN/m³. The thickened parameters in the table are the values of other influencing factors in the single factor study. For example, when the elastic modulus of the roadway floor is taken as a variable, the buried depth of the roadway is 500 m, the thickness of the floor is 2.5 m, the width of the roadway is 5.0 m and the widths of the fracture zone are 2.0 m and 1.0 m, respectively. The variation curve of the displacement of the bottom plate with the influence factors is shown in Figure 7.

It can be seen from Figure 7 that (1) when each influencing factor is taken as a single variable, the displacement of the floor is approximately symmetrically distributed along the middle position of the floor and the displacement of the floor decreases from the middle to both sides and (2) when the floor position is fixed, the displacement of the floor is negatively correlated with the elastic modulus and the thickness of floor, and positively correlated with the buried depth of roadway, the width of roadway and the width of fracture zone. An analysis shows that, the larger the elastic modulus and floor thickness, the stronger the resistance to deformation of the floor beam and the higher the stability of the floor. The greater the buried depth of the roadway, the greater the load borne by floor beam and the greater the displacement of floor. As the width of roadway and the width of fracture zone increase, the free surface of the roadway becomes larger, and the larger the range of load releasing floor agglomeration, the larger the floor displacement.

To sum up, the generation mechanism of floor heave in roadways can be divided into two steps. First, the floor beam of the roadway generates an upward displacement under the comprehensive action of overlying rock load and lower floor load transmitted by coal pillars on both sides of roadway. Then, the floor beam accelerates the deformation under horizontal stress. As the compressive strength of the floor beam is much greater than its tensile strength, the strength of the outer surface of the floor beam decreases due to the effect of tensile stress in the deformation process. In addition, the rock beam is a heterogeneous material, so random cracks occur on the outer surface of the rock beam and the integrity decreases, resulting in floor heave under the comprehensive action of the lower floor load and horizontal load. In order to reduce floor displacement, the displacement of the floor beam can be reduced by increasing the elastic modulus and thickness of the floor beam and reducing the width of the roadway and the width of the fracture zone. Specific field measures are as follows:

(1) Measures to increase the elastic modulus of floor beam: floor grouting reinforcement and construction of an inverted bottom arch. Grouting reinforcement can increase the strength and bending stiffness of the bottom plate. The construction of inverted arch can lead to the advantages of a strong compressive ability of the outer surface of the inverted arch device, and transfer the load transmitted by the overlying strata to deep rock mass through the inverted arch device, so as to protect the stability of the floor.

(2) Measures to increase the thickness of floor beam: floor construction anchor and grouting reinforcement. A floor bolt can anchor the stratified rock beam of the floor as a whole to form a combined floor, increase the overall thickness and strength of the floor beam and weaken the stratification effect of the floor. Grouting reinforcement can improve the strength of floor, increase the cohesion between adjacent rock strata and improve the overall strength of the floor beam.

(3) Measures to reduce the width of the roadway and the fracture zone: grouting reinforcement at the side and construction of anchor bolts at the bottom corner. Grouting reinforcement of the side surrounding rock can increase the overall stability of the side surrounding rock, reduce the width of the fracture zone of the side wall surrounding rock and reduce the actual span of the roadway floor rock beam and the width of the free surface of the rock beam. A construction floor angle anchor bolt can anchor the upper layered rock beam and the lower rock layer into a whole, shorten the span of the rock beam and increase the stability of the floor.

3. Stability Analysis of Roadway Floor

3.1. Model Settings

In order to further study the influence characteristics of different factors on the stability of roadway floor, a three-dimensional finite difference numerical simulation software (FLAC^3D 5.0) was used to analyze the evolution of floor displacement. The size of the model is 300 m × 225 m × 50 m (length × width × height), and the Mohr–Coulomb constitutive model was used as the material failure criterion. In order to simplify the calculations of the model, half of the working face is selected to establish the model according to the principle of symmetry. In the model, the working face is arranged along the y-direction, with a length of 150 m, and is advanced along the x-direction. The size of the roadway section for transportation in the working face is 5.4 m × 3.6 m, the size of roadway section for auxiliary transportation is 4.6 m × 3.8 m and the width of coal pillar in the section is 35.0 m. In order to eliminate the influence of the boundary effect, 30.0 m boundary coal pillars are set on the left and right sides of the model in the x-direction and 30.0 m boundary coal pillars in the y-direction. The numerical model is established as shown in Figure 8.

In order to reduce the amount of calculations, the model is divided into four areas along the advancing direction of the working face, and the advancing distance of the working face in each area is 60 m. The first area is set as the original geological condition as the control group. Area 2, area 3 and area 4 study the influence on floor displacement by enhancing the elastic modulus of floor, strengthening the strength of surrounding rock and increasing the thickness of floor strata, respectively. Strengthening the elastic modulus of the floor and strengthening the strength of the side wall rock were realized by setting the relevant parameters of the floor and the side wall rock. For increasing the thickness of the floor strata, considering that the mudstone is divided into two layers (each layer is 1.3 m thick) in the model, the interface command of the software is used to simulate the bonding degree of the two layers of mudstone. The contact surface parameter was smaller in excavation areas 1 to area 3, and the contact surface parameter was larger in excavation area 4. The plane distribution of the model is shown in Figure 9.

The vertical and horizontal displacements were fixed at the bottom of the model and the horizontal displacement was limited around the model. According to the buried depth of the coal seam and the modeling height, a uniform load of 15.0 MPa was applied above the model to simulate the weight of the overlying strata. The physical and mechanical parameters of each stratum in the model are shown in

Table 2. Physical and mechanical parameters of rock.

Lithology	Bulk Modulus (GPa)	Shear Modulus (GPa)	Cohesion (MPa)	Tensile Strength (MPa)	Density (kg·m−3)	Internal Friction Angle (°)
Grit	7.35	6.63	3.56	4.35	2720	40
Mudstone	3.57	1.10	2.20	1.90	2400	30
Coal	2.06	1.01	2.98	1.45	1321	25
Fine Sandstone	5.19	3.26	2.91	2.83	2640	43
Siltstone	4.30	1.87	5.37	3.73	2350	36
Sandstone	3.61	1.48	2.80	2.80	2360	35

.

3.2. Evolution Characteristics of Floor Surface Displacement

In combination with the spatial position (30–34.6 m in the y-direction) and layout direction of the working face auxiliary transportation roadway in the model, the vertical displacement nephogram of the roadway floor surface is drawn as shown in Figure 10.

It can be seen from Figure 10 that the floor displacement in the four areas gradually decreases from the middle of the roadway to the left and right sides, and the overall displacement of the roadway floor in the four areas is the largest in Area 3, followed by Area 1 and Area 2, and the smallest in Area 4.

In order to clearly show the size of and change in floor displacement in the four study areas, the floor displacement data of the roadway were extracted and the floor displacement distribution curves of different study areas of the roadway were established, as shown in Figure 10. The legend Y = 31.0 represents the displacement distribution curve of the roadway floor at 31.0 m in the y-direction of the model, and Y = 32.3 and Y = 33.6 are the same.

It can be seen from Figure 11 that:

(1) The displacement variation in the roadway floor in different research areas is large, but the overall trend in the changes is basically the same. The displacement in the middle of the roadway (Y = 32.3) is greater those that on both sides of the roadway (Y = 31.0 and Y = 33.6).

(2) By comparing the floor displacement distribution curves of the four study areas, the overall floor displacement decreases in the order area 3 > area 1 > area 2 > area 4; that is, from the perspective of controlling floor displacement, the effect of increasing the thickness of floor strata is the best, followed by enhancing the elastic modulus of floor strata and finally strengthening the strength of surrounding rock.

(3) The displacement of the floor in the third area is greater than that in the first area. On analysis, it can see that the third area studies the influence characteristics of the strength of the surrounding rock on the roadway floor displacement. In the model, by increasing the parameters of the coal body on both sides of the corresponding area, the strength difference between the surrounding rock on the side and the floor is large and the roadway floor is used as an unloading area because of its small strength. As a result, the floor displacement of the three positions in the area increases.

In order to verify the correctness of the above conclusion, the displacement data of the side of the auxiliary haulage roadway were extracted, and the displacement distribution curves of the surrounding rock in the side of different study areas of the roadway were established, as shown in Figure 12. The legend Y30-Z18.5 represents the displacement distribution curve of the roadway side at the intersection of 30.0 m in the y-direction and 18.5 m in the z-direction in the model, and the other legends are the same. At the same time, it is stipulated that the displacement of the left side of the roadway is positive and the displacement of the right side is negative.

It can be seen from Figure 12 that the displacement change trend in the roadway sides in different study areas is consistent, but the displacement change characteristics of the left and right sides are quite different. The overall displacement of the left side of the roadway is greater than that of the right side of the roadway, which is caused by the fact that the right side is close to the working face. With the advance of the working face, the overlying strata bend and sink as a whole and are supported by the floor strata. The position of the section coal pillar belongs to the pressure relief area, and the force of horizontal stress on the coal pillar is reduced, while the left side is in a state of concentrated stress, so the displacement of the right wall of the roadway is smaller. At the same time, the overall displacement of different research areas on the left side of the roadway decreases in the order area 1 > area 2 > area 3 > area 4. The overall displacement of different study areas on the right side of the roadway decreases in the order of area 4 > area 2 > area 1 > area 3. The strength parameters of the surrounding rock in the third area are strengthened; therefore, the displacement of the surrounding rock in the third area is reduced, and its displacement variation characteristics explain that the displacement of the floor in the three areas is greater than that in the first area.

3.3. Evolution Characteristics of Deep Floor Displacement

In order to further reveal the influence of various influencing factors on the stability of the floor, the vertical displacement cloud map of the deep position of the floor (Z = 16.3 m plane in the model) was drawn, and is shown in Figure 13.

It can be seen from Figure 13 that the overall distribution trend of deep floor displacement is consistent with the roadway surface displacement nephogram, and the floor displacement gradually decreases from the middle of the roadway to the left and right sides. The overall displacement of the four study areas decreases in the order area 3 > area 1 > area 2 > area 4.

In order to clarify the displacement change law of the deep floor strata, the vertical displacements of the different depths of the roadway floor at the center of the four study areas were extracted in combination with the occurrence characteristics of the roadway floor strata and the division of the floor strata by the four study areas in the numerical model, taking into account the grid size in the model. The data extraction position is shown in Figure 14.

According to the extracted displacement data, the vertical displacement distribution curves of different positions away from the roadway surface were established. At the same time, the displacement change rate, R, was defined in order to clearly reflect the degree of synergistic deformation between deep strata. See Formula (1) for the calculation of displacement change rate:

$$R=\frac{S_{i+1}-S_i}{S_i}$$

(9)

where S_i and S_i+1 are the vertical displacement of two adjacent measuring points and S_i+1 is the vertical displacement of the measuring point near the roadway floor.

Distribution curves of vertical displacement and displacement change rates at different positions away from the roadway surface were established and are shown in Figure 15. In the legend, 60 m, 120 m, 180 m and 240 m are the data of Lines 1 to 4, representing the displacement change curve of different depths of the floor from Area 1 to Area 4. A change rate of 60 m represents the displacement change rate of two adjacent measuring points on Line 1, and the others are the same.

It can be seen from Figure 15 that:

(1) The distribution curves of displacement and displacement change rate of the roadway floor at different depths in each research area have the same trend.

(2) For the displacement of the floor at different depths, the displacement gradually decreases with the increase in the distance from the roadway surface. The displacement decreases greatly within 2.6 m from the floor surface and becomes smaller after 2.6 m, where the displacement basically remains unchanged. Area 3 has the largest displacement in the same position of the floor, followed by area 1 and area 2, and then area 4 has the smallest displacement. In other words, area 4 has a better control effect on the floor surrounding rock. By analyzing the causes of the above phenomena, it can be seen that the immediate bottom of the roadway is mudstone with an average thickness of 2.6 m, while the lower part of the mudstone is thick hard sandstone. The strength difference between the two rocks is large, so the floor displacement varies greatly within a depth of 2.6 m.

(3) For the displacement change rate of the floor at different depths, as the distance from the roadway surface increases, the displacement change first increases, then decreases and then gradually remains unchanged. The displacement change rate reaches a maximum value at a floor depth of 1.73 m and rapidly decreases within the floor depth range of 1.73~2.6 m. When the depth of floor exceeds 2.6 m, the displacement change rate decreases slightly. Additionally, the displacement change rate of the roadway floor at the same position decreases in the order area 4 > area 1 > area 2 > area 3. Compared with area 1, the cooperative deformation ability of the floor strata in area 2 and area 3 is enhanced; that is, increasing the elastic modulus of floor strata and strengthening the surrounding rock can effectively control the deep displacement of the floor. The change rate of displacement in area 4 is relatively high. It can be seen from the analysis that the mudstone is divided into two layers in the numerical model, and the contact surface command of the software is used to simulate the bonding degree of the two layers of mudstone. The existence of the contact surface destroys the grid connection between the two layers of mudstone and changes the cooperative deformation capacity of the rock stratum. Although the change rate of the floor displacement is large, the displacement at the same depth of the floor is the smallest; that is, increasing the thickness of the floor strata has the best effect on controlling the floor heave of the roadway.

4. Discussion and Application of the Results

The deformation and failure process of the surrounding rock of roadways is the result of comprehensive action of many factors. Different geological conditions and roadway service conditions cause different characteristics of surrounding rock deformation and failure. The 301 working face auxiliary transportation roadway has a large burial depth, a low floor mudstone strength and is easy to soften when exposed to water. Meanwhile, due to the long period of influence of the dynamic mining pressure of the working face and the lack of support measures for the floor, the roadway floor heave is serious.

4.1. Discussion of Results

The theoretical analysis results reveal that floor heave occurs under the action of initial upward displacement and later accelerated deformation. Based on this, the floor heave can be controlled by increasing the elastic modulus and thickness of roadway floor strata and reducing the depth, width and width of the roadway fracture zone. However, due to the existing situation of coal seam and the minimum requirements for roadway section size, the buried depth, horizon and section size of the roadway are difficult to change. Therefore, the elastic modulus and thickness of the floor strata should be increased, e.g., by implementing floor grouting reinforcement and constructing anchor bolts, to increase the overall thickness and strength of the floor rock beam and weaken the floor layering effect.

The numerical simulation results verify the reliability and accuracy of the theoretical analysis. By analyzing the evolution characteristics of surface displacement and deep displacement of the roadway, it is determined that increasing the thickness of floor strata is the best way to control floor heave, followed by enhancing the elastic modulus of the floor strata and finally by strengthening the side surrounding rock

4.2. Field Application

The secondary deformation of the roadway floor was controlled by the method of “floor lifting + bottom angle bolt + floor bolt” in the auxiliary transportation roadway of the 301 working face in a mine after roadway floor heave. The specific method comprises the following steps: excavating and cleaning the floor heave part of the auxiliary transportation roadway of the 301 working face when the normal use of the roadway is affected due to large floor heave, installing floor angle anchor bolts at appropriate positions of floor angles on two sides of the roadway (if the original floor angle anchor rods are not loosened and damaged, the floor angle anchor bolts should be installed at positions near the original floor angle bolts) and finally constructing three anchor bolts at the floor of the roadway. The row spacing of the anchors was 1500 mm × 1600 mm. Φ22 × 2800 mm metal anchor bolts were used for the floor angle anchor bolts and the floor anchor bolts, and one MSK23/35 and two MSZ23/60 resin anchoring agents were used for a single anchor bolt. The reinforcement support of an auxiliary haulage roadway of the working face 301 is shown in Figure 16.

After the floor heave was treated by the method of “floor lifting + bottom angle bolt + floor bolt” in the auxiliary transportation roadway of the 301 working face, the roadway had a small floor heave and the floor deformation was about 300 mm at most during the service of return air roadway of the 303 working face. Adopting the method of “floor lifting + bottom angle bolt + floor bolt” is equivalent to increasing the overall thickness of floor strata, which has a good effect on controlling the floor heave, and indirectly verifies the results of the theoretical analysis and numerical simulations.

5. Conclusions

(1) The main factors affecting the floor heave of the auxiliary transportation roadway in the 301 working face are determined as follows: the buried depth of the roadway is large, the period of dynamic pressure is long, the characteristics of surrounding rock are unique and the support strength is insufficient.

(2) A mechanical model of the roadway floor rock beam is established, and the floor displacement is quantified when each load acts alone. It is revealed that the floor heave is caused by an initial upward displacement and later by accelerated deformation. It is also determined that the floor displacement is negatively correlated with elastic modulus and floor thickness and positively correlated with buried depth, roadway width and fracture zone width.

(3) A numerical calculation method is used to study the characteristics of the floor displacement by enhancing the floor elastic modulus, strengthening the surrounding rock and increasing the thickness of the floor rock. Through the analysis of floor surface displacement and deep displacement, it is determined that increasing the thickness of the floor rock is the best way to control floor heave, followed by enhancing the floor elastic modulus and then by strengthening the surrounding rock.