Research on the integrated technology of bearing structure reconstruction and support control in high risk area of top coal roadway in thick coal seam

Abstract

In order to address the problems of easy breakage of surrounding rock, large roof subsidence and frequent roof falling and rib spalling in the high roof area and structural belt of the top-coal caving roadway in thick coal seam, the research on the surrounding rock control technology of the top-coal caving roadway in thick coal seam is carried out by means of field investigation, laboratory test, theoretical analysis, numerical simulation and industrial-scale test, taking the grout transport crossheading of I040905 working face of a coal mine as the engineering background. The geological characteristics and deformation and failure characteristics of the roadway were clarified through field investigation, and the physical and mechanical parameters of the surrounding rock were determined by laboratory tests. Based on theoretical analysis, the mechanical model of the roof of the top-coal caving roadway is established, and the deformation and failure mechanism of roof and roadway side and the synergistic deformation law of roof and side are revealed. The influence of top coal thickness and strength on the stability of surrounding rock is analyzed by FLAC3D. The combined control technology of ' advanced duct pre-grouting and bolt-cable combined support ' is proposed, and the support parameters are optimized. The test results show that technology can effectively control the deformation of the roadway. The maximum subsidence of the roof is 111.6 mm, and the maximum displacement of the two ribs is 85.3 mm. The force of the bolt (cable) is uniform and stable, which greatly improves the efficiency of roadway excavation. It successfully addresses the problem of safe excavation in the high roof area and structural belt of the top-coal caving roadway in thick coal seam, providing a theoretical basis and practical engineering reference for similar roadways. The research results will also contribute to human safety and environmental protection in coal mining operations.

Introduction

China’s coal resources are abundant, and the reserves of thick coal seams with a thickness of more than 3.5 m account for about 40% −60% of the total coal reserves in China. In the past decade, new mines have mostly taken thick coal seams as priority mining objects. The development level of thick coal seam mining technology has become a key indicator to measure the progress of China ‘s coal science and technology and the overall strength of the industry^1,2,3. The mechanized mining technology with large mining height has been widely concerned and applied in the mining of thick coal seams because of its outstanding advantages such as high yield and efficiency, high resource recovery rate and low production cost^4,5. It has become the mainstream mining technology of 4.0–7.0 m thick coal seams in China. The corresponding top-coal caving roadway section is generally large, and the top coal and the rib coal are integrated coal bodies, with low strength, poor integrity, and large ground stress in superimposed deep mining. It leads to frequent accidents of roof separation, two sides bulging and roof falling when the roadway is too high in the roof area or structural belt⁶. According to statistics, since 2017, roof accidents are still high-risk accidents in coal mines⁷. The annual repair and maintenance costs of the top-coal roadway in thick coal seams are huge, which seriously restricts the safe production of mines.

Domestic and foreign scholars have carried out a lot of research on the stability of top-coal caving roadway^8,9,10,11. Xiao et al.¹² revealed the influence of mining depth, section size and tectonic stress on the stability of top-coal caving roadway using numerical simulation. Bai et al.¹³ used UDEC-Trigon model to study the displacement field, stress field and plastic zone distribution law of the top-coal caving roadway in thick coal seam. The results showed that the core reason for the shear failure of the roadway roof was that the distribution of shear stress and friction stress at the rib-roof corner angle of the two s ribs of the roadway was not balanced. Tian et al.¹⁴ carried out a drilling hole inspection on the roof of the top-coal caving roadway in the thick coal seam and found that the reason for the instability of the roof of the top-coal caving roadway was that there were many weak surfaces, which reduced the overall strength of the top coal. Chai et al.¹⁵ analyzed the bearing deformation and failure characteristics of large-span top-coal open-off cut top-coal, and found that the caving zone above the top-coal was asymmetrical failure pattern. Wang et al.¹⁶ put the layered and nonlinear failure characteristics of rock roof into the design of bolt support, and put forward a simplified design method. Su et al.¹⁷ put forward the coupling support concept of ' composite beam-compression bearing arch ‘, and optimized the support bearing structure of roadway with extra-thick composite roof. However, the existing research mostly ignores the characteristics of ' side-to-top synergistic deformation ' of top coal in thick coal seam, and the reconstruction technology of bearing structure for high roof area and tectonic belt is insufficient, which is difficult to meet the site requirement.

Therefore, considering factors such as research representativeness, engineering suitability and field conditions, the roadway of I040905 working face in a coal mine is selected as the site of the project. The working face is highly consistent with the geological conditions of the top-coal caving roadway focused on in the research. There are not only typical key working conditions such as the thickness and strength difference of top coal and the fluctuation of lateral pressure coefficient, but also the roadway excavation process and the occurrence characteristics of surrounding rock (such as coal cohesion, internal friction angle and other mechanical parameter characteristics) are generally common with similar top-coal caving roadways. The research results can be directly applied to similar engineering scenarios. In order to systematically carry out research, the deformation characteristics of the roadway in the working face were investigated in detail through field investigation, and the mechanical parameters of the surrounding rock were accurately obtained by combining the indoor test. In the process of research, through the combination of theoretical analysis and numerical simulation, the core influencing factors of the stability of the top-coal caving roadway are further revealed, and the reconstruction technology of the bearing structure is put forward. Finally, through the field test in the working face, the feasibility and reliability of the proposed technology are fully verified. The research conclusions can provide an important reference for the surrounding rock control engineering of similar top-coal caving roadways in China.

Engineering geological conditions and mechanical properties of surrounding rock

Engineering background

The roadway of I040905 working face in a coal mine is selected as the project site of this study. The grouting transportation crossheading of the working face is connected to the I040905 working face cut in the north and the transportation lane in the south. The total design length is 2473.93 m, which mainly undertakes the coal transportation and personnel passage function of the working face. The main mining No.3 coal seam is located in the middle and lower part of the Shanxi Formation, and the whole mine field can be mined. According to the statistical data of No.131 and M11 boreholes in the working face, the average thickness of the coal seam is 6.06 m (Fig. 1).

Fig. 1

Mining engineering plan for working face.

The lithology of the immediate roof of the coal seam no. 3 of the I040905 working face is mainly sandy mudstone, and the rock is a dark gray block structure. The strength of the rock mass is semi-hard, and the corresponding roof stability category is moderately stable. The main roof is mainly composed of fine-grained sandstone, and the local area is intercalated with thin-layer siltstone, which belongs to the hard rock layer. The direct bottom is dominated by sandy mudstone, and the rock is black sandy argillaceous, which belongs to semi-hard rock. It is easy to expand in water and is a moderately stable roof. The old bottom is dominated by fine-grained sandstone, and the rock is grayish white. The comprehensive histogram is shown in Fig. 2.

Fig. 2

I040905 working face coal seam roof and floor lithology comprehensive histogram.

Determination of physical and mechanical parameters of surrounding rock

For laboratory tests, rock samples were obtained by drilling and coring at positions below 39 m of the roof of the grouting and transportation gateway in the I040905 working face. Specimens were processed strictly in accordance with MT/T 38‑2019 which is a standard method for the determination of physical and mechanical properties of coal and rock. The RMT‑150B rock mechanics test system was used to determine parameters including uniaxial compressive strength and tensile strength, while an UTA2000A ultrasonic detector was employed to measure acoustic wave velocity. The test results are presented in Table 1.

Table 1 Test results of rock physical and mechanical parameters.

These experimentally derived physical and mechanical parameters were directly integrated into subsequent theoretical and numerical analyses. The elastic modulus, Poisson’s ratio, and uniaxial compressive strength were adopted as core inputs to define the homogeneous elastic material properties of the top coal and coal mass in the elastic mechanics-based roof beam model, enabling the quantification of stress distribution and deformation characteristics The full suite of parameters from Table 1, including cohesion, internal friction angle, and deformation modulus, were assigned to corresponding rock and coal strata in the FLAC3D model to calibrate the Mohr-Coulomb constitutive model. This allowed the simulation to accurately reproduce surrounding rock mechanical response under varying top-coal conditions and predict roadway stability and deformation.

Analysis of mine pressure characteristics of roadway

In view of the problems of roof stress concentration near the fault zone of the grout transport crossheading in the I040905 working face, the top coal is broken and the roof is easy to leak, as shown in Fig. 3. The existing grouting, pipe shed, and other methods are not effective due to the lack of development of coal cracks and the difficulty of slurry penetration, resulting in difficult support, slow driving speed and high safety risk. Therefore, it is necessary to study and determine the effective reinforcement method of top coal broken section and optimize the support parameters, to ensure the safe and rapid excavation and long-term safe use of the roadway.

Fig. 3

Destruction and deformation of the roof of the grout transport trough.

Deformation and failure mechanism of the surrounding rock in thick coal seam top-coal caving roadway

Mechanical model of the roof and deformation characteristics

Combined with the theoretical analysis method, based on the theory of elastic mechanics, the mechanical model of the roof of the top - coal caving roadway is established. The top coal and coal body are included in the model analysis category. The top coal and coal body are regarded as homogeneous elastic materials, and the roof is simplified into a simple supported beam and equal section beam combination model, which establishes a theoretical and data foundation for the subsequent in-depth study of the stability of roadway surrounding rock^18,19.

Fig. 4

Top-coal structural mechanic’s model.

According to the simply supported beam model, the top coal is subjected to the uniform load q of the overlying strata, and the differential equation of the deflection curve is²⁰:

$$\frac{{{d^2}\omega }}{{{d^2}{x^2}}}= - \frac{{M\left( x \right)}}{{El}},$$

(1)

In the Eq. (1), E is the elastic modulus of top coal; l is the cross-section moment of inertia of top coal, and ${M_{(x)}}$ is the function of bending moment of top coal, $M\left( x \right)=\frac{1}{2}qax - \frac{1}{2}q{x^2}$; q is the weight of top coal; a is the width of the roadway.

Equation (1) continuous integral:

$$\frac{{d\omega }}{{dx}}= - \frac{1}{{El}}\left( {\frac{{qa}}{4}{x^2} - \frac{q}{6}{x^3}+{C_1}} \right),$$

(2)

$$\omega = - \frac{1}{{El}}\left( {\frac{{qa}}{{12}}{x^3} - \frac{q}{{24}}{x^4}+{C_1}x+{C_2}} \right),$$

(3)

From the symmetry, ${\omega _{\left( 0 \right)}}={\omega _{\left( a \right)}}=0$, $\omega _{{\left( 0 \right)}}^{\prime }=0$. Substituting into Eqs. (2) and (3), we can get ${C_1}= - \frac{{q{a^3}}}{{24}}$, ${C_2}=0$. Then the differential equation of the deflection curve can be solved as:

$$\omega = - \frac{1}{{El}}\left( {\frac{{qa}}{{12}}{x^3}+\frac{q}{{24}}{x^4} - \frac{{q{a^3}}}{{24}}x} \right)=\frac{{qx}}{{24El}}\left( {{x^3} - 2a{x^2}+{a^3}} \right),$$

(4)

It can be derived that the maximum subsidence position of the roof is the midpoint of the roadway, and the maximum subsidence value is:

$${\omega _{\hbox{max} }}={\omega _{\left( {\frac{a}{2}} \right)}}=\frac{{5q{a^4}}}{{384EI}},$$

(5)

It can be seen from Eq. (5) that the maximum subsidence position of the roof is the midpoint of the roadway, and the top coal caving phenomenon will occur when the roof subsidence exceeds the failure limit. The roof subsidence of the top-coal caving roadway is proportional to the fourth power of the roadway width, which is inversely proportional to the elastic modulus and the cross-sectional moment of inertia of the top coal. Since, the roadway failure is closely related to the roadway width and the top coal strength, therefore, in the process of roadway excavation, it is necessary to strictly control the problem of ultra-wide and ultra-high roadway caused by over-excavation, to avoid the increase of roadway roof width and the low strength of top coal, which leads to roof breaking and even falling.

The thickness of the top coal in the top-coal caving roadway of the I040905 working face of a coal mine is 2.5 m, the width of the roadway is 5.1 m, and the span-height ratio is calculated to be 2.04 m. This parameter is close to the definition standard of the simply supported single-span deep beam. The shear stress formula of the cross-section beam is introduced to calculate the roof deflection of the top - coal caving roadway. The maximum deflection formula is^21,22:

$${\omega _{\hbox{max} }}\omega \left( {\frac{a}{2}} \right)=\frac{{5q{a^4}}}{{384EI}}\left[ {1+\frac{{48\left( {1+\mu } \right){h^2}}}{{25{a^2}}}} \right]$$

(6)

In the Eq. (6), $\frac{{5q{a^4}}}{{384EI}}$ is the maximum deflection solution obtained by the calculation method of simply supported beam. Equation (6) shows that, compared with the calculation results of simply supported beams, the calculation results of uniform cross-section beams increase the shear deformation of simply supported single-span deep beams under uniform load. In this study, according to the specific working conditions of the I040905 working face of a coal mine, the top coal load q = 0.17 MPa, the top coal thickness h = 2.5 m, the elastic modulus E = 2.36GPa, the Poisson ‘s ratio µ = 0.31, the roadway width a = 5.1m, and the cross-section moment of inertia I = h³/12 were substituted in Eq. (6). By analyzing the correlation characteristics and calculation results between the variables, it is revealed that the variation law of roof deflection (roof subsidence) is positively correlated with roof load and Poisson’s ratio and negatively correlated with top coal thickness and elastic modulus, and approximately exponentially correlated with roadway width. It is obvious that the key of controlling the roof subsidence of the top-coal caving roadway is the reasonable control of the roadway width, and the increase of the roadway width will bring about an exponential growth change.

Roadway deformation mechanics analysis

Displacement analysis of roadway rib

After the excavation of the top-coal caving roadway, when the physical and mechanical properties of the coal body in the roadway rib are weak, the coal body in the coal rib failure zone and the plastic strain zone is in the limit equilibrium state. Because the Poisson ‘s ratio of the coal body is larger than that of the rock mass above the top coal, and the cohesion and internal friction angle of the coal body at the junction of the roof and floor are lower than the cohesion and internal friction angle of the coal body in the rib, the coal body in the limit equilibrium area has a tendency to squeeze out to the inside of the roadway under the action of the roof and floor strata, which will produce a horizontal normal shear stress ${\tau _{xy}}$on the interface between the coal body in the rib and the top coal^23,24. The calculation model of the interface stress of the coal rib is shown in Fig. 5. In the figure, P represents the support resistance of the support structure applied to the coal rib, h is the height of the coal rib, ${x_0}$ represents the width of the stress limit equilibrium zone, and ${\sigma _x}$is the average value of the horizontal stress along the thickness direction of the whole coal seam at $x={x_0}$.

Fig. 5

Coal interface stress calculation model of roadway rib.

The total displacement S of roadway rib is derived by elastic-plastic theory¹⁷:

$$S=\frac{{ - \lambda k\gamma H}}{{E\sqrt \beta }}\left[ {2coth\left( {\sqrt \beta {x_0}} \right) - csch\left( {\sqrt \beta {x_0}} \right)} \right] - \frac{{\Delta m}}{{m - \Delta m}}{x_0},$$

(7)

In the Eq. (7); $\beta ={K_s}/(mE)$ (K_s is the tangential stiffness of the interface between the coal body the roof and floor); k is the stress concentration factor; λ is the lateral pressure coefficient on the elastic-plastic interface; E is the elastic modulus of coal; m is the height of the unit.

From the analysis of Eq. (7), it is derived that the deformation of the coal body in the roadway rib is positively correlated with the lateral pressure coefficient, the peak stress of the overlying strata and the width of the limit equilibrium zone of the coal body, and negatively correlated with the elastic modulus of the coal body (i.e., the comprehensive strength of the coal). At the same time, there is a significant simultaneous evolution effect between the deformation process of the roadway rib and the roof subsidence phenomenon. The deformation mechanism of ' sidewall-roof integration ' of the surrounding rock in top-coal roadway is revealed. Therefore, its support design must follow the principle of collaborative control of the roof.

Analysis of the limit equilibrium zone of the roadway

$${x_0}=\frac{{\lambda h}}{{2\tan {\varphi _0}}}\ln \left( {\frac{{k\gamma H+\frac{c}{{\tan \varphi }}}}{{\frac{c}{{\tan \varphi }}+\frac{P}{\lambda }}}} \right)$$

(8)

After the completion of the roadway excavation, the stress field of the surrounding rock will be redistributed, and the coal s rib will be destroyed under the action of stress disturbance, and the damage range will gradually expand to the deep part of the coal body, thus forming three areas of deformed zone, plastic zone and elastic zone²⁵. The part of the coal body from the deformed zone, the plastic zone to the elastic zone boundary is in the stress limit equilibrium state as a whole. Based on the limit equilibrium theory, the Equation for calculating the width of the limit equilibrium zone ${x}_{0}$ can be derived as²⁶:

In the formula, ${x}_{0}$ represents the width of the limit equilibrium zone; $h$ is the height of the roadway; k represents the stress concentration factor corresponding to the vertical stress peak position of the overlying strata of the roadway; γ is the average bulk density of the overlying strata of the roadway; H refers to the actual overburden depth of the roadway; λ represents the lateral pressure coefficient of this region; P is the support resistance of the support structure acting on the coal rib; c and φ correspond to the cohesion and internal friction angle of the coal body itself, respectively. It has been derived from Eq. (8) that:

• (1) The key to ensure the stability of the roadway is whether the length of the bolt can pass through the limit equilibrium zone. When the bolt passes through the limit equilibrium zone and has sufficient length to anchor in the elastic zone, the inward movement and deformation of the roadway can be well controlled.

• (2) The higher the support resistance of the support structure to the roadway rib is, the smaller the width of the limit equilibrium zone of the roadway s rib is. Therefore, the higher the prestress of the bolt cable in the anchorage range can increase the stability of the coal rib of the roadway^27,28,29.

• (3) Due to the integrity of the top coal and the coal in the top-coal caving roadway, when the width of the limit equilibrium zone of the coal rib of the roadway is increasing, the damage degree of the coal rib will be further extensive, which is equivalent to the mechanical properties and strength of the surrounding rock of the two ribs of the roadway. Both are reduced, resulting in an increase in the equivalent width of the roadway, a decrease in the elastic modulus of the coal body, and an increase in the deformation degree of the top coal. Therefore, in the support of the top-coal caving roadway, the support of the roadway is also significantly important.

Analysis of influencing factors of roadway stability

The top-coal caving roadway of I040905 working face is affected by the bending and vertical stress superposition of the side coal causing the bending deformation to occur. The top coal is broken, and the stress is concentrated. The roof coal rock beam is prone to separation or synergistic deformation. Its stability is closely related to the factors such as roadway span and support mode, and the larger the roadway span, the greater the tensile stress borne by the roof coal rock beam, and the worse the stability is. The high prestressed support can inhibit the roof subsidence, the large anchorage zone can form an effective composite beam structure, and the prestressed diffusion can reduce the load of coal rock beam. At the same time, the top coal and the coal side are integrated, and the instability of the coal side will equivalently increase the roadway span and initiate the roof failure. Based on the above deformation mechanism and stress analysis, the roof support of the top-coal caving roadway should follow the five principles of high strength, high prestress diffusion, large anchorage zone, reliability of supporting structure and simultaneous treatment of side and roof, in order to ensure the stability and safety of roadway excavation.

Numerical simulations

Numerical model establishment

In order to further explore the influence of top coal thickness and strength on the stability of surrounding rock, this study uses numerical simulation method to construct a rectangular section of top-coal caving roadway model with FLAC3D numerical simulation software. The section size of the roadway is designed to be 5.1 m × 3.5 m, which is arranged along the coal seam floor. The overall size of the model is 65.1 m long, 10 m wide and 53.9 m high. The Mohr-Coulomb constitutive model is used for coal and rock materials (Fig. 6). This numerical simulation focuses on the two key factors of top coal thickness and strength, and systematically analyzes its influence on the stability of roadway surrounding rock.

Fig. 6

Numerical model diagram of top coal roadway.

Simulation program

Each scheme only changes one parameter to study the influence of various factors on the stability of the top-coal caving roadway. Among them, the coal and rock material parameters are set as follows: the bulk modulus is 2.07 GPa, the shear modules is 0.90 GPa, the tensile strength is 0.188 MPa, the cohesion is 4.35 MPa, the internal friction angle is 25.2 °, and the density is 1293 kg/m³. The strength of top coal is defined as E. When adjusting the thickness of top coal, only the bulk modulus and shear modulus are multiplied by the corresponding multiples. The simulation scheme is shown in Table 3.

Table 3 Numerical simulation scheme.

Analysis of the key affecting factors

The influence of different top coal thickness on the deformation, stress distribution and plastic zone development of roadway surrounding rock

In order to explore the influence of different top coal thickness on the deformation law of roadway surrounding rock, the vertical displacement cloud diagram and horizontal displacement cloud diagram of roadway under different top coal thickness conditions are obtained by simulation, as shown in Figs. 7 and 8.

Fig. 7

Vertical displacement distribution map of roadway surrounding rock under different top coal thickness.

Fig. 8

Horizontal displacement distribution map of roadway surrounding rock under different top coal thickness.

The simulation results of vertical displacement show that (Fig. 7), when the thickness of top coal increases, the position of the roadway will move down relatively, which leads to the change of roadway excavation mode from top excavation to bottom excavation. Since the strength of coal body is lower than that of direct roof and floor, the amount of floor heave will increase with the decrease of coal body strength, while the amount of roof subsidence will decrease with the increase of roof strength. The horizontal displacement distribution shows that (Fig. 8), the change of top coal thickness has little effect on the horizontal displacement of roadway surrounding rock, and there will be no large fluctuation. The reason is that the change of the roadway layer does not change the nature of the coal body in the roadway, so the horizontal displacement of the two ribs does not change much.

In order to study the influence of different top coal thicknesses on the stress evolution law of roadway surrounding rock, the vertical stress cloud diagram and horizontal stress cloud diagram of roadway under different top coal thicknesses are obtained by simulation, as shown in Figs. 9 and 10.

Fig. 9

Vertical stress distribution map of roadway surrounding rock under different top coal thicknesses.

Fig. 10

Horizontal stress distribution map of roadway surrounding rock under different top coal thicknesses.

The simulation results of vertical stress show that (Fig. 9), the ' anti-butterfly airfoil ' stress concentration area is formed on both ribs of the roadway, and the concentration area migrates from the top angle to the bottom angle with the increase of top coal thickness. The horizontal stress distribution shows that (Fig. 10), the thickening of the top coal reduces the range of the horizontal stress concentration area of the roof and expands the stress reduction area, indicating that the stress state changes from compression to tension. At the same time, the stress concentration of the floor is triggered by the thinning of the coal body.

In order to study the influence of different top coal thickness on the expansion law of plastic zone of roadway surrounding rock, the distribution cloud diagram of plastic zone of roadway surrounding rock under different top coal thicknesses is obtained by simulation, as shown in Fig. 11.

Fig. 11

Distribution of plastic zone of roadway surrounding rock under different top coal thicknesses.

Figure 11 show that the distribution of plastic zone of roadway surrounding rock under different top coal thicknesses. The plastic zone first appears interlayer failure in the roof of the roadway, and then interlayer failure occurs in the floor. The butterfly plastic zone of the roadway rib gradually expands to the deep of the two sides at the top corner of the roadway.

Fig. 12

Variation trend diagram under different top coal thicknesses.

Figure 12 show that the variation trend diagram under different top coal thicknesses. By comparing and analyzing the evolution law of roadway displacement, stress and the expansion characteristics of plastic zone under different top coal thickness, the key factors of surrounding rock stability of top-coal caving roadway with different top coal thicknesses are systematically studied. The results show that the deformation of the roof will increase with the increase of the thickness of the top coal, the amount of floor heave decreases with the increase of the thickness of the top coal, and the rib of the roadway is not greatly affected. Therefore, attention should be paid to the selection of reasonable layout layers to control the deformation of roadways during roadway excavation.

The influence of different top coal strength on the deformation, stress distribution and plastic zone development of roadway surrounding rock

In order to study the influence of different top coal strength on the deformation law of roadway surrounding rock, the vertical displacement cloud diagram and horizontal displacement cloud diagram of roadway under different top coal strengths are obtained by simulation, as shown in Figs. 13 and 14.

Fig. 13

Vertical displacement distribution of roadway surrounding rock under different top coal strengths.

Fig. 14

Horizontal displacement distribution map of roadway surrounding rock under different top coal strengths.

The simulation results show that (Figs. 13 and 14), the increase of top coal strength can effectively inhibit the deformation of roadway surrounding rock: the roof subsidence and the convergence of the two ribs are significantly reduced, while the floor deformation is slightly affected. The quantitative relationship between the displacement of surrounding rock and the strength of top coal is shown in Fig. 17, which confirms that improving the strength of top coal is the key approach to control the stability of roof and two ribs.

In order to study the influence of different top coal strengths on the stress evolution law of roadway surrounding rock, the vertical stress cloud diagram and horizontal stress cloud diagram of roadway under different top coal strengths are obtained by simulation, as shown in Figs. 15 and 16.

Fig. 15

Vertical stress distribution map of roadway surrounding rock under different top coal strengths.

Fig. 16

Horizontal stress distribution map of the roadway surrounding rock under different top coal strengths.

The simulation results (Figs. 15 and 16) show that the change of top coal strength has little effect on the vertical and horizontal stress distribution of roadway surrounding rock. In terms of vertical stress, only the range of stress concentration area at the bottom corner of roadway increases slightly with the increase of strength. The horizontal stress field is basically unaffected.

In order to study the influence of different top coal strengths on the expansion law of plastic zone of the roadway surrounding rock, the distribution cloud diagram of the plastic zone of the roadway surrounding rock under different top coal strengths is obtained by simulation, as shown in Fig. 17.

Fig. 17

Plastic zone distribution map of roadway surrounding rock under different top coal strengths.

The simulation results show that (Fig. 17), the increase of top coal strength can significantly improve the failure mode of surrounding rock: the plastic zone of roof and two ribs is changed from large-scale tensile-shear mixed failure to shear failure, and the tensile failure zone of floor is also significantly reduced. This shows that improving the strength of top coal can effectively enhance the tensile capacity of coal and inhibit cracking, therefore comprehensively improve the overall stability of the roadway surrounding rock (Fig. 18).

Fig. 18

Variation trend under different top coal strengths.

By comparing and analyzing the displacement stress evolution and plastic zone expansion of roadway under different top coal strengths, the influencing factors of surrounding rock stability of top-coal caving roadway with different top coal strengths are studied. The strength of the top coal has little effect on the stress and plastic zone. It mainly reduces the deformation of the roadway by controlling the strength of the roadway roof. It can be found that the higher the strength of the top coal, the smaller the roof deformation and the smaller the displacement of the rib. The characteristics of the combined deformation of the top-coal caving roadway are further verified.

Numerical simulation of support effect of advanced pipe pre-grouting combined control technology in thick coal seam top-coal of the roadway

The establishment of model and the determination of simulation scheme

Based on the numerical model of the top coal of the roadway established above, a 30 m extension is made in the Y direction. The roof of the 10 m roadway is the area of easy caving, and the parameters of the area of easy caving are shown in Table 4. The remaining 20 m roadway is the normal section. The numerical model is shown in Fig. 19. The size of the model is 65.1 m long, 40 m wide and 53.9 m high. The model has 410800 grids and 429680 nodes. The section of the roadway is a rectangular section of 5.1 m × 3.5 m, and the Mohr-Coulomb model is used. Figure 20 shows the layout of the advanced ducts in each simulation scheme.

Table 4 Mechanical parameters of rock mass in easy caving area.

Fig. 19

Slice diagram of top coal of the roadway model in easy caving area of thick coal seam.

Fig. 20

The schematic diagram of the advanced duct arrangement scheme of each simulation scheme. (a)Original support, ༈b༉Plan A, ༈c༉Plan B, ༈d༉Plan C.

Analysis of deformation and failure characteristics of roadway under different schemes

Based on the above model, four schemes are used to excavate and support the roadway respectively, and the supporting effect of different schemes on the easy caving area of the roadway is compared horizontally, and the displacement cloud map under different supporting schemes is extracted, as shown in Fig. 21.

Fig. 21

Comparative contour map of roadway surrounding rock deformation under different support schemes.

In order to quantitatively compare the original scheme and the three optimization schemes, the displacement of the same monitoring point in different schemes is monitored and the data is extracted in the numerical calculation process, (Table 5). The displacement of the center point of the roof deformation is the value of the roof subsidence. The maximum displacement point of the floor deformation is the value of the floor heave. The maximum horizontal displacement point of the two ribs is the displacement of the rib part. The maximum vertical displacement point of the left and right top angles of the roof is the subsidence of the top angle, and the supporting effect of different schemes is compared horizontally.

Table 5 Comparison of deformation of monitoring points in caving section.

Based on the analysis of the effects of the three support schemes, it can be concluded that although only the roof grouting reinforcement of the top-coal caving roadway is carried out, it also has a strong control effect on the deformation of the roadway side. Controlling the subsidence of the roof is conducive to reducing the squeezing deformation of the roadway side due to the vertical load caused by the subsidence of the roof, indicating that the deformation of the top-coal caving roadway is a phenomenon of coordinated deformation of the side and the top.

Fig. 22

Comparison of control effects of different support schemes on roadway surrounding rock.

The numerical simulation results show that (Fig. 22), the grouting reinforcement of the roof can effectively control the deformation of the surrounding rock of the top coal of roadway, and the roof subsidence decreases with the increase of the layout density of the advanced pipe, but its strengthening effect tends to decrease after the pipe spacing is less than 500 mm. The control effect of the first scheme (the highest degree of roof strengthening) is the best. Not only the deformation of the roof and the top angle is the smallest, but also the two ribs are significantly inhibited by the rib-roof combined effect, which confirms the key role of high-strength roof reinforcement in maintaining the overall stability of the roadway.

In order to further compare the supporting effect of each scheme on the top coal of roadway, the plastic zone distribution of the four schemes is shown in Fig. 23.

Fig. 23

The failure of roadway surrounding rock under different support schemes is compared with the contour diagram of plastic zone.

The numerical simulation results show that (Fig. 23), the plastic zone range of the optimization scheme is smaller than that of the original support. By observing the plastic zone range of the rib, it can be seen that although the three schemes only carry out grouting reinforcement on the top of the roof and the coal body of the rib, they can still greatly reduce the plastic zone expansion of the rib. It is derived that the deformation of the top - coal caving roadway is the process of combined influence of the rib and the top. The reduction effect of the three schemes on the expansion range of the plastic zone of the floor is not obvious. If the floor heave area is strong, the floor needs to be treated.

Although the control effect of Scheme 1 is technically optimal, it needs to arrange more dense advanced pipes. Compared with Scheme 2, the additional reduction of surrounding rock deformation caused by Scheme 1 is only about 5%, with low marginal benefit and lower economical benefit. The second scheme can reduce the roof subsidence by about 79.54% and the convergence of the two ribs by about 57.30%, while effectively ensuring the safety of the roadway and reasonably reducing the maintenance cost. In summary, Scheme 2 is determined to be the best choice for both technical and economic benefits.

Field test

In order to improve the physical and mechanical properties of the roof, enhance the integrity of the roof, and control the damage and deformation of the surrounding rock of the top coal of the roadway, this study proposes a combined control technology of advanced catheter pre-grouting, and adopts the method of industrial-scale test to carry out the technical application verification in the field of grouting transportation in I040905 working face, so as to evaluate the feasibility and control effect of this technology in engineering practice.

This technology adopts anchor net cable support, advanced 2-inch pipe (steel), grouting (quick-setting double liquid material), in which the advanced pipe adopts 2-inch pipe, the diameter of grouting borehole is 50–75 mm, the spacing of borehole is 500 mm, and the tail of cotton yarn winding pipe is used to seal the hole. The length of sealing hole is about 2000 mm, then the pipe is grouted, and the grouting pressure is more than 15 MPa. After the slurry penetrates and fills the inner and outer space, a catheter like concrete with quick-setting materials inside and outside is formed, which plays a supporting role in the roof and can effectively prevent the roof from falling^30,31,32,33.

In view of the fact that the selected double-liquid grouting material has the characteristics of rapid setting and early strength (15 min strength ≥ 11 MPa, 40 min strength ≥ 24 MPa), the excavation and support process can be carried out immediately after the grouting operation. This process can increase the tunneling footage of 3–5 rows per shift and 6–10 m per day, which is 3–5 times higher than the daily footage of 1–2 m in the original high caving area, and can always maintain a lead reinforcement distance of not less than 1000 mm. The daily cycle execution of the installation catheter → grouting → excavation support process can ensure the safe and efficient passage of the roadway through the top coal caving area. The layout scheme of the lead pipe is shown in Fig. 24.

Fig. 24

Advanced catheter schematic diagram.

The order of roadway reinforcement is: drilling advanced pipe hole → installing advanced pipe → grouting to the broken surrounding rock → next stage roadway excavation.

In order to ensure the safety and efficiency of roadway excavation, the mine pressure observation was carried out during the excavation of I040905 conveyor belt transport crossheading in a coal mine, and the law of mine pressure observation during excavation was validated. Two mine pressure observation stations (#1 ~ and #2) were set up within 200 m of changing the support parameters, and the spacing was 100 m. The first station was set up at 50 m away from changing the support parameters^34,35.

The cross-point method is used to monitor the displacement of the preset displacement monitoring section. The first measurement is carried out on the day of the layout of the test section, and then the observation is carried out every 3 days. The obtained displacement curve of the roadway surface is shown in Fig. 25.

Fig. 25

Roadway surface displacement monitoring curve.

The field application shows that (Fig. 25), the combined control technology of advanced catheter pre-grouting for top coal of roadway proposed in this paper can effectively solve the problem of top coal caving and super elevation in the easy caving area. The deformation of the roadway tends to be stable for 36 days after excavation. Finally, the convergence of the two ribs is 79.5–85.3 mm, the convergence of the roof and floor is 105.2–111.6 mm, and the amount of floor heave is very small (about 4 mm), which confirms that the technology has excellent results in controlling the stability of surrounding rock of such roadways.

The roof displacement monitor is used to monitor the separation of the surrounding rock of the roadway. One roof displacement monitor is installed on each measuring surface, and two stations. The first measurement was carried out on the day of the test section arrangement and then observed every 3 days. The two base points of the roof displacement monitor are located at 2.5 m and 8 m above the roadway roof. The monitoring results are shown in Fig. 26.

Fig. 26

The internal separation amount of roadway roof at each measuring point.

The mine pressure monitoring shows that the deep separation measurement of the roadway roof is maintained at 0.3–1.5 mm after the optimized support, and the shallow separation measurement is 1.7–4.2 mm, which is much lower than the critical value of 20 mm. The bed separation mainly occurs in the rock layer of 2.5–8 m above the top coal, and it is basically stable within 30 days after excavation, which confirms that the support scheme can effectively control the deformation of the surrounding rock of I040905 grout transport crossheading.

The second bolt above the two ribs of the roadway and the bolts at the two corners were selected to monitor the stresses of the bolts. The bolts from the left side of the roadway were numbered 1 to 4 clockwise; the anchor cable selects an anchor cable in the middle of the roadway roof, numbered 1. The measurement frequency is once on the day after installation and then measured once every 3 days. The results obtained by summarizing the observation data of each measuring point are shown in Fig. 27.

Fig. 27

The change trend diagram of bolt cable force in each section roadway.

The monitoring results show that (Fig. 27), the stress of the bolt and the anchor cable in the optimized support system increases simultaneously and tends to be stable, which is highly correlated with the displacement curve of the surrounding rock, confirming its combined bearing mechanism. Compared with the original scheme, the average prestresses of the roof and rib bolts after optimization is reduced to 171 kN and 142.5 kN, respectively, and the average prestress of the anchor cable is reduced to 286 kN, with a significant decrease. This phenomenon shows that the new scheme fully stimulates the self-bearing capacity of the surrounding rock by means of roof grouting and advanced pipe, and achieves long-term stability under lower support load, and the action of support is excellent.

Conclusion

Aiming at the problem of surrounding rock control in thick coal seam top-coal roadway, taking the grout transport crossheading of I040905 working face of a coal mine as the engineering study site, through systematic research methods such as field investigation, laboratory test, theoretical modeling, numerical simulation and field test, the top- rib combined deformation mechanism of top-coal roadway and its application path in support design are developed, the instability law of surrounding rock is revealed, the targeted control technology is proposed, the applicable boundary of the results is defined, and the research limitations and future directions are analyzed. The main research results are as follows:

• (1) The instability of the surrounding rock of the top coal roadway is controlled by the fault structure and the coordinated deformation of the top-side. The stress concentration near the fault structure zone is easy to cause the surrounding rock to be broken. The traditional support scheme does not consider the coupling effect of the top-rib deformation, and it is difficult to form an effective overall bearing structure only by implementing local support. It is easy to cause instability phenomena such as top coal caving, roadway super elevation and two- ribs bulging, which eventually leads to support failure.

• (2) The thickness and strength of top coal are the main controlling factors affecting the stability of surrounding rock: the increase of top coal thickness is easy to initiate the subsidence of roof and floor, and then squeeze the surrounding rock of the side to enlarge the deformation of the side; the strength of top coal is negatively correlated with the deformation of surrounding rock. The lower the strength of top coal, the weaker the bearing capacity of roof is, the more significant the combined deformation effect of top and rib, and the worse the stability of surrounding rock.

• (3) Advanced catheter pre-grouting + bolt-cable ' collaborative support technology can effectively control the deformation of surrounding rock in top-coal roadway. Advanced catheter pre-grouting can significantly improve the physical and mechanical properties of top-coal and surrounding rock, and enhance the integrity and collaborative bearing capacity of top-wall surrounding rock. Bolt-cable support can construct the integrated bearing structure of top-wall and constrain the development of top-wall deformation. The field test shows that after the application of this technology, the maximum subsidence of the roof is 111.6 mm, and the maximum displacement of the two ribs is 85.3 mm. The deformation of the surrounding rock is controlled within the allowable range of the specification, and the supporting effect is good.

• (4) The collaborative support technology can effectively address the problem of safe and rapid excavation of top-coal roadway in thick coal seam. Its core design ideas and technical parameters can provide technical reference and engineering reference for surrounding rock control engineering of top-coal roadway under similar geological conditions, and have good widespread adoption and application value.