1 Introduction

With increased depletion of shallow coal resources, coal mining is gradually developing into deep seams, facing high stress, ultra-soft rock, intense mining activities, and other complex geological conditions (Gao et al. 2021; Jiao et al. 2021; Kang et al. 2014; Konicek and Waclawik 2018; Tan et al. 2021; Xie et al. 2020; Yu et al. 2020). At present, the gob-side entry driving method with coal pillars has been widely used in deep coal mining. In this method, two mining roadways need to be excavated and one coal pillar needs to be reserved for mining a working face (Bai et al. 2015; Wang et al. 2021a, b; Zha et al. 2017). A huge gob is formed behind the working face after coal mining in which the gob roof strata collapse to compress the area of coal pillar. Under the aforementioned complex conditions, different degrees of stress concentration exist in the coal pillars (He et al. 2021; Wang et al. 2020, 2022), the coal body interior at the working face (Feng et al. 2016), and the rock mass of the roof (Lu et al. 2015) in the deep mining roadways. However, the shallow and weak rock mass of the roadway roof is often loose and broken due to closing to the roadway excavation unloading face, which is challenging to support and reinforce (Li et al. 2018; Li et al. 2019; Wang et al. 2021b; Jiang et al. 2023). This uneven distribution of stress often leads to accidents such as large deformation in rocks, roof collapse, and rock bursts, significantly increasing the difficulty of roadway safety control (Wei et al. 2018).

Aiming at addressing stress concentration in coal and rock masses in deep mine roadways, scholars have carried out relevant research on pressure relief of coal and rock masses (Chen et al. 2021; Kang et al. 2018; Wei et al. 2011; Yu 2016; Zhang et al. 2018). In terms of coal pressure relief at the working face, Zhang et al. (2018) conducted laboratory and field tests of drilling pressure relief, and optimized parameters to effectively reduce stress concentration in coal seams. Chen et al. (2021) performed simulations of water injection in a coal seam ahead of the working face and found that water injection can significantly reduce the peak stress of the coal seam. In terms of rock pressure relief in the roadway, Kang et al. (2018) carried out hydraulic fracturing field tests for pressure relief, which effectively reduced the rock stress in deep mining roadway roof. Wei et al. (2011) conducted theoretical analysis and field tests on parameter optimization of roadway roof blasting, and the optimized parameters effectively improved the pressure relief effect of the in situ rock mass.

The aforementioned research on coal and rock pressure relief has effectively promoted the development of pressure relief technology and made significant contributions to ensure the safe mining of coal mines. However, these studies did not consider the stress concentration in roadway pillars, but rather focused on the local pressure relief of roadway surrounding rock (Yang et al. 2016; Zhang et al. 2020). Limited study exists on the comprehensive pressure relief technology of rock mass that changes the stress transfer structure of roadway roof. To solve this problem, our research group proposed the method of automatically formed roadway through roof cutting and pressure relief (Wang et al. 2018). This method has been successfully applied in shallow mines. The automatic formation of the gangue rib by mining pressure and expansion characteristics of rock masses enables the formation of roadways and the elimination of coal pillars. Moreover, the roof pre-splitting technology cuts off the stress transfer between the gob and the roadway roof. However, this method does not consider the pressure relief in the coal body at the working face, and the shallow rock mass of deep mining roadway is mostly weak and broken. The control method of considering comprehensive pressure relief in coal and rock masses in deep mining roadways and ensuring the safety of support and reinforcement system needs to be further studied.

In this work, we take the Suncun Coal Mine, the deepest mine in China, as the engineering background to analyze the deformation and failure behavior of surrounding rock in gob-side entry driving with coal pillars. We carry out field tests and numerical simulation on the control mechanism of high-stress deep mining roadway. We study the variation laws of surrounding rock stress and displacement under different rock pressure relief and coal pressure relief parameters, and clarify the control mechanism of surrounding rock pressure relief and anchor grouting reinforcement of the roadway. Subsequently, we propose an integrated control method of comprehensive pressure relief for coal and rock masses and anchor grouting reinforcement in deep mining roadways. Compared to other control methods in deep mining roadways, the characteristics of this method are as follows. In order to solve the problem of “different degrees of stress concentration in the coal pillars, the coal body interior at working face and rock at roadway roof deep position” and “broken rock at roadway roof shallow position”, it utilizes roof directional pre-splitting and coal body drilling to achieve comprehensive pressure relief in different directions. Meanwhile, it utilizes high prestressed grouting to improve the self-supporting capacity of broken surrounding rock. Lastly, we conduct geomechanical model tests and field application of the new method to verify the correctness of the numerical simulation and the rationality of the new method.

2 Deformation and failure analysis of the high-stress deep mining roadway

2.1 Engineering background

Suncun Coal Mine is located east of the Xinwen coalfield in China with a maximum mining depth of 1500 m, which is a typical deep high-stress coal mine. The altitude of working face 2223 is − 1238.9 to − 1303.2 m. Its average dip length is 261.7 m, average strike length is 569.2 m, and the average coal thickness is 2.9 m, as shown in Fig. 1. The immediate roof is siltstone, with an average thickness of 2.5 m and well-developed bedding and a compressive strength of 7.7 MPa. The basic roof is sandstone, with an average thickness of 6.5 m and a compressive strength of 38.4 MPa. The immediate floor is fine sandstone, with an average thickness of 3.0 m and a compressive strength of 19.0 MPa.

Fig. 1
figure 1

Working face histogram and roadway reinforcement parameters

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The traditional method of gob-side entry driving with coal pillars is used in the haulage roadway of the working face, and coal pillar of 5 m is reserved. In terms of support and reinforcement parameters, traditional anchor mesh support is applied to the roof of the roadway. The roof cable is Φ 22 mm × 5000 mm, and the roof bolt is Φ 20 mm × 2800 mm, with a spacing of 1000 mm × 800 mm. In terms of pressure relief parameters, the drilling pressure relief is applied in the coal body, with a drilling depth of 15 m, a drilling diameter of 150 mm, and a spacing of 5 m. No pressure relief measures have been taken for the rock mass. The specific parameters of roadway support and pressure relief are shown in Fig. 1.

2.2 Monitoring and analysis of roadway surrounding rock

Under the above parameters of roadway support and pressure relief, the haulage roadway surrounding rock at the working face 2223 is seriously deformed, as shown in Fig. 2. In order to clarify the reason for surrounding rock deformation and bolt failure of this type of roadway, we select a typical section to monitor and analyze the deformation, loosening range, and reinforcement system stress. Figure 3 shows the evolution of surrounding rock deformation and bolt (cable) stress during coal mining. The middle position of the roadway roof, left shoulder, right shoulder, solid coal side, and coal pillar side are recorded by a borehole televiewer. The hole depths are 8 m, 8 m, 8 m, 8 m, and 5 m, respectively, at the five locations mentioned above. According to the distribution of fractures in the holes, the surrounding rock within the detection range is divided into a deep stable zone, a middle microfracture zone, and a shallow fracture zone, as shown in Fig. 4.

Fig. 2
figure 2

Roadway deformation and failure in the coal mine

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Fig. 3
figure 3

Evolution of surrounding rock deformation and bolt (cable) force as a function of distance from working face

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Fig. 4
figure 4

Zoning of roadway surrounding rock

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According to the analysis in Figs. 2, 3 and 4:

  1. (1)

    During the coal mining period, the rock deformation increases as the distance from the working face reduces. Specifically, under the influence of strong excavation of the working face, the rock deformation rate increases sharply from 71 to 0 m ahead of the working face. Deformation is characterized by convergence deformation of the two sides greater than roof and floor convergence deformation. The maximum deformation of the roadway surrounding rock occurs near the shoulders on both sides, with a value up to 689 mm.

  2. (2)

    The integrity of deep rock mass is relatively high, and the loose fracture range of shallow rock mass is large on the roadway. Overall, it exhibits the loose fracture characteristics of “rock mass of the roadway roof > rock mass of left shoulder > coal pillar side > rock mass of right shoulder > solid coal side”. Among them, the maximum shallow fracture zone of the roadway roof reaches 3.5 m, and there is almost no crack development in coal and rock masses 5.5 m outside the surrounding rock.

  3. (3)

    The stress of the roadway roof bolts shows a trend of “first increasing, then decreasing” during the coal mining period. The bolts are affected by the expansion of the shallow fracture zone, the anchor section cannot be in close contact with the surrounding rock, and the stress is significantly reduced within 20 m in front of the working face. The stress of roadway roof cable always shows an increasing trend, with a maximum stress of 388.5 kN, which is close to the ultimate breaking force of this type of cable.

To sum up, the reasons for the difficulty of controlling such roadways are as follows. The shallow rock mass of the roadway roof is very soft with developed bedding, and the compressive strength is only 7.7 MPa. The shallow rock mass is loose and broken severely under the superposition of high in situ stress and strong mining-induced stress during mining. Because the whole bolt is placed in the shallow fracture zone, the reinforcement effect is poor, so it is difficult to restrain the expansion of the loosening range. At this point, due to the existence of roadway pillars, the stress in the deep rock mass of the roadway roof and solid coal side is concentrated. As the shallow fracture zone expands, if the surrounding rock stress is released slowly, large deformation will be expected. If the surrounding rock stress is released instantaneously, roof fall, rock bursts and other dynamic disasters can occur. Therefore, the key to solving the above problems lies in the integrated control of reinforcement and pressure relief for the deep and shallow coal and rock mass in the roadway.

3 Numerical simulation of high-stress deep roadway safety control mechanism

To clarify the rock control mechanism of this high-stress deep mining roadway, and to explore a reasonable pressure relief and reinforcement mode, we conduct numerical simulation on comprehensive pressure relief in coal and rock mass and anchor grouting reinforcement.

3.1 Anchor grouting parameters by field test

We select a typical section in the haulage roadway of the working face 2223 to carry out field testing of anchor grouting parameters in order to obtain the reinforcement range of grout diffusion and the degree of improvement in rock strength, which is the foundation for the development of numerical simulation.

Based on the original support, the grouting cable and dynamometer cable are added. To ensure that the grouting cable is anchored to stable rock stratum above 1.0 m, the model of the supplementary grouting cable is Φ 22 mm × 6000 mm. The model of the supplementary dynamometer cable is Φ 22 mm × 5000 mm. The spacing between adjacent dynamometer cables is 400 mm, numbered I1–I4, as shown in Fig. 5a. The drawing force of the supplementary dynamometer cable is tested, compared with the cable without grouting reinforcement, the improvement of the drawing force is evident, as shown in Fig. 5b. In addition, borehole televiewer is also used to monitor grout diffusion before the installation of dynamometer cables, as shown in Fig. 5c.

Fig. 5
figure 5

Field test of anchor grouting parameters

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According to the analysis of Fig. 5a–c:

  1. (1)

    Grout is clearly distributed in the range of 0–1200 mm from the grouting cable as 6000 mm of grouting cable is used to reinforce the roadway roof. The result shows that the effective grout diffusion range is 0–1200 mm under this working condition, and the average grout diffusion height in this range is 2800 mm.

  2. (2)

    Compared with the non-grouting area cable, the drawing force of dynamometer cable I1–I3 is increased to 137.5%, 130.0%, and 112.5%, respectively, with an average of 126.7% in the grout diffusion area. The drawing force of the dynamometer cable I4 is basically the same as that of non-grouting area cable. These data also verify that the effective grout diffusion ranges from 0 to 1200 mm.

3.2 Numerical scheme design

The roadway section size and reinforcing parameters of roadway side are kept the same. The numerical simulation scheme consistent with field pressure relief parameters is set, numbered N. On this basis, we perform four types of comparisons, i.e., cutting height, cutting angle, drilling spacing, and drilling length, represented by I, II, III, and IV, respectively. The specific design parameters are given in Table 1. In the above four schemes, the shallow rock mass of the roadway roof is strengthened by improving the mechanical parameters of the surrounding rock mimic grouting effect. According to the field test results of anchor grouting parameters, the mechanical parameters of surrounding rocks are increased by 1.27 times within the range of 0–2800 mm of the roadway roof.

Table 1 Design table of numerical simulation scheme
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3.3 Model establishment and evaluation index

  1. (1)

    Establishment of the numerical model

The numerical model is established by using the 3DEC numerical software and the size is 200 m × 80 m × 50 m (width × height × thickness), as shown in Fig. 6. The lower boundary is fixed, the horizontal displacements on the side boundaries are restricted, and the upper boundary is applied with a stress boundary. The uniformly distributed load is applied to simulate the overburden, so that the horizontal stress at the working face of the model is 23.4 MPa and the vertical stress is 19.4 MPa. The Mohr Coulomb failure criterion is used, and its physical and mechanical parameters are consistent with the field conditions of the coal mine. Furthermore, a comprehensive monitoring section is arranged in the middle of the model body to monitor the stress and displacement distribution of the rock (within 0–30 m) in schemes N and I–IV respectively. The number of the shoulder monitoring points is recorded as S-k, where k is from 1 to 15.

  1. (2)

    Establishment of evaluation index

Fig. 6
figure 6

Numerical calculation model

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To better analyze the variation of rock stress and displacement in the roadway in different schemes, quantitative evaluation indexes such as stress reduction rate of shulder \(\upeta _{{{\text{S}}_{{{\text{ij}} - {\text{k}}}} }}\) and deformation control rate of the surrounding rock \(\updelta _{{{\text{ij}}}}\) are established.

  1. (a)

    Stress reduction rate of shoulder \(\upeta _{{{\text{S}}_{{{\text{ij}} - {\text{k}}}} }}\)

$$\upeta _{{{\text{S}}_{{{\text{ij}} - {\text{k}}}} }} { = }\frac{{{\text{S}}_{{{\text{N}} - {\text{k}}}} - {\text{S}}_{{{\text{ij}} - {\text{k}}}} }}{{{\text{S}}_{{{\text{N}} - {\text{k}}}} }} \times 100\%$$
(1)

where \(\upeta _{{{\text{S}}_{{{\text{ij}} - {\text{k}}}} }}\) is the stress reduction rate of the k monitoring point in the roadway shoulder in scheme ij, which represents the degree of stress reduction at the k monitoring point in the roadway shoulder in scheme ij relative to the stress at the corresponding monitoring point in the original scheme N. Higher \(\upeta _{{{\text{S}}_{{{\text{ij}} - {\text{k}}}} }}\) means better pressure relief effect. i represents different scheme numbers, including the four schemes designed (scheme I–IV). j represents the sub-scheme for different schemes, and j is from 1 to 5.

\({\text{S}}_{{{\text{N}} - {\text{k}}}}\)—Stress at the k monitoring point of the roadway shoulder in the original scheme N, MPa.

\({\text{S}}_{{{\text{ij}} - {\text{k}}}}\)—Stress at the k monitoring point of the roadway shoulder in scheme ij, MPa.

  1. (b)

    Deformation control rate of the surrounding rock \(\updelta _{{{\text{ij}}}}\)

$$\updelta _{{{\text{ij}}}} { = }\frac{{{\text{D}}_{{\text{N}}} - {\text{D}}_{{{\text{ij}}}} }}{{{\text{D}}_{{\text{N}}} }} \times 100\%$$
(2)

where \(\updelta _{{{\text{ij}}}}\) is the deformation control rate of the surrounding rock in scheme ij, which represents the reduction degree of the roof and floor convergence deformation in scheme ij relative to that in the original scheme N. Higher \(\updelta _{{{\text{ij}}}}\) means better control effect. i and j have the same meaning as Eq. (1). A negative value indicates that the roof and floor convergence deformation in this scheme is higher than that in the original scheme N. A positive value means that the roof and floor convergence deformation in this scheme is lower than that in the original scheme N.

\({\text{D}}_{{\text{N}}}\)—The roof and floor convergence deformation of the roadway in the original scheme N, mm.

\({\text{D}}_{{{\text{ij}}}}\)—The roof and floor convergence deformation of the roadway in scheme ij, mm.

3.4 Analysis of test results

  1. (1)

    Results analysis of scheme I

According to the analysis in Fig. 7:

  1. (a)

    Roadway shoulder stress: The shoulder stress of the roadway in the direction away from the roadway surface, in the original scheme (scheme N) and different roof cutting height schemes (scheme I1–I5), exhibits the trend of “first increasing, then decreasing and finally stabilizing”, as shown in Fig. 7a. Peak stress on the shoulder appears 4 m from the surface of the roadway and has a value of 59.4 MPa in scheme N. Peak shoulder stress occurs 6 m from the roadway surface and is reduced in varying degrees in schemes I1–I5. The stress reduction rates are 2.8%, 5.6%, 17.6%, 17.2%, and 9.6%, respectively, as shown in Fig. 7b. The results show that the directional roof pre-splitting can reduce the rock stress in the deep mining roadway, transfer the peak stress of the shoulder to the deep formation, and the pressure relief effect first increases and then decreases with increasing roof cutting height.

  2. (b)

    Roadway surrounding rock deformation: The roof and floor convergence deformation in the original scheme (scheme N) is 688 mm, higher than that in each roof cutting height scheme (scheme I1–I5). The deformation control rates of surrounding rock \(\updelta _{{{\text{ij}}}}\) in different roof cutting height schemes are 29.5%, 44.0%, 48.8%, 21.5%, and 4.1%, respectively. As shown in Fig. 7b, the control effect of the roadway surrounding rock first increases and then decreases with increasing roof cutting height, as shown in Fig. 7b. The reason for this is that the rock mass after roof cutting is broken and expanded to fill the gob. With the increase of the cutting height, the range of broken and expanded rock mass increases, which gradually weakens the influence of the roof rotation and the gob subsidence. However, as the volume of expanded rock mass within the range of roof cutting height is exactly equal to the volume of the gob, the roof cutting height continues to increase, and the pressure relief and control effect of the roadway surrounding rock decreases.

Fig. 7
figure 7

Evolution of stress and displacement of the roadway surrounding rock in scheme I

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  1. (2)

    Results analysis of scheme II

According to the analysis in Fig. 8:

  1. (a)

    Roadway shoulder stress: Roadway shoulder stress exhibits a similar trend of “first increasing, then decreasing, and finally stabilizing”, in the direction away from the roadway surface, in the original scheme (scheme N) and different roof cutting angle schemes (scheme II1–II5), as shown in Fig. 8a. Peak stress on the roadway shoulder appears 4 m from the roadway surface with a value of 59.4 MPa in scheme N. Peak shoulder stress occurs 6 m away from the roadway surface, but in scheme I5, peak shoulder stress appears 4 m away from the roadway surface. The shoulder stress is reduced to varying degrees in each roof cutting angle scheme, with reduction rates of 4.5%, 16.7%, 17.6%, 10.5%, and 1.5% respectively, as shown in Fig. 7b. The results show that the directional roof pre-splitting can reduce the rock stress in the deep mining roadway, transfer the peak stress of the shoulder to the deep formation, and the pressure relief effect first increases and then decreases with increasing roof cutting angle.

  2. (b)

    Roadway surrounding rock deformation: The roof and floor convergence deformation in the original scheme (scheme N) is 688 mm, higher than that in cutting height schemes I1–I5. The deformation control rates \(\updelta _{{{\text{ij}}}}\) in roof cutting angle schemes are 34.5%, 51.3%, 48.8%, 31.8%, and 9.2%, respectively. As shown in Fig. 7b, the control effect increases first and then decreases as the roof cutting angle increases. The reasons for this are as follows. The rock mass after roof cutting is broken and expanded to fill the gob due to rock expansion characteristics. As roof cutting angle increases, the rock mass within the cutting range is prone to collapse under the action of dead weight, thus weakening the impact of gob roof collapse on the surrounding rock of the roadway. However, if the roof cutting angle is too large, the cantilever beam of the roadway roof is too long, the broken and expansion characteristics of the rock cannot be fully utilized to fill the gob, so the pressure relief and control effect of the roadway surrounding rock is reduced.

  3. (c)

    Results analysis of scheme III.

Fig. 8
figure 8

Evolution of stress and displacement of the roadway surrounding rock in scheme II

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According to the analysis in Fig. 9:

  1. (a)

    Roadway shoulder stress: Similarly, the roadway shoulder stress shows a “first increasing, then decreasing, and finally stabilizing” trend, in the direction away from the roadway surface in different pressure relief drilling spacing schemes (schemes III1–III5), as shown in Fig. 9a. With increasing drilling spacing (scheme III1–scheme III5), shoulder stress gradually increases and the stress reduction rate gradually decreases. The stress reduction rate decreases slowly within the drilling spacing range of 1–5 m (scheme III1–scheme III3), and it decreases significantly within the range of 5–9 m (scheme III3–scheme III5). The results show that the pressure relief drilling spacing can be shortened, which can effectively reduce the surrounding rock stress under roof cutting and pressure relief. However, when drilling spacing is reduced to a certain extent, the pressure relief effect is not obvious, and the construction cost and difficulty are improved.

  2. (b)

    Roadway surrounding rock deformation: When the pressure relief drilling spacings are 1 m (scheme III1) and 3 m (scheme III2), the convergence deformation of roof and floor is significantly higher than that of other schemes (scheme III3–III5), at 1150 mm and 1030 mm, respectively. As the pressure relief drilling spacing increases to 5 m, the convergence deformation of the roof and floor is significantly reduced to 352 mm. The results show that the pressure relief drilling spacing should not be reduced too much for roof directional pre-splitting, so as to prevent the deformation, instability and failure of the roadway surrounding rock caused by excessive pressure relief.

  3. (c)

    Results analysis of scheme IV

Fig. 9
figure 9

Evolution of stress and displacement of the roadway surrounding rock in scheme III

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According to the analysis in Fig. 10:

  1. (a)

    Roadway shoulder stress: The stress of roadway shoulder shows the similar trend of “first increasing, then decreasing, and finally stabilizing” in the direction away from the roadway surface in different pressure relief drilling length schemes (schemes IV1–IV5), as shown in Fig. 10a. The shoulder stress reduction rates for all schemes are 8.8%, 10.8%, 17.6%, 19.1%, and 20.2%, respectively. The results show that the longer the pressure relief drilling length, the better the relief effect, especially as the pressure relief drilling length is increased from 10 to 15 m.

  2. (b)

    Roadway surrounding rock deformation: As the pressure relief drilling length increases, the deformation control rate of the surrounding rock \({\updelta }_{{{\text{ij}}}}\) increases first and then decreases. When the drilling lengths are 15 m (scheme IV3) and 20 m (scheme IV4), the surrounding rock control effect is the best, and the deformation control rates are 48.8% and 50.2%, respectively, as shown in Fig. 10b. The results show that there should be a suitable range of pressure relief drilling length. Too short length can lead to insufficient pressure relief in the coal body, and too long length can cause excessive pressure relief in the coal body, resulting in deformation, instability, and rock failure.

Fig. 10
figure 10

Evolution of stress and displacement of the roadway surrounding rock in scheme IV

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3.5 Summary

  1. (1)

    Anchor grouting reinforcement technology with high prestressed grouting cable as the core can effectively improve the integrity of the shallow broken roadway roof and fully realize its self-supporting capacity. Taking the working condition in this study as an example, the average height of grout diffusion is 2800 mm, the effective range of grout diffusion is 0–1200 mm, and the drawing force of the supporting system in the grouting reinforcement range is increased by 126.7%.

  2. (2)

    Roof pre-splitting in deep roadway can effectively reduce the stress concentration in the surrounding rock. As the critical parameter of the roof pre-splitting, the height and angle of roof cutting should have a reasonable design range. When the roof cutting height is too short or the roof cutting angle is too large, the rock that collapses into the gob cannot be completely cut from the top of the roadway. In addition, the broken and expanded rock to fill the gob is insufficient, and the pressure relief effect is not satisfactory.

  3. (3)

    Drilling pressure relief for coal can effectively release the internal stress of the surrounding rock. As the critical parameter of drilling pressure relief, the drilling spacing and length of pressure relief should all have reasonable design ranges. Too small drilling spacing or too large drilling length can lead to excessive pressure relief in the coal body, which is not conducive to the stability of the roadway surrounding rock. If drilling spacing is too large or drilling length is too small, the pressure relief will be insufficient, and then the stress accumulated in the coal body cannot be effectively released. Taking the working condition in this work as an example, the reasonable drilling spacing is about 5 m, and the proper drilling length is 15–20 m.

4 Verification of the comprehensive pressure relief control method and anchor grouting reinforcement

4.1 Control concept and method

Based on the results of field testing and numerical simulation results, we propose a control method of comprehensive pressure relief for coal and rock and anchor grouting reinforcement in deep mining roadways (“CPR-AG method”), as shown in Fig. 11. We expect this method could solve the problem of “different degrees of stress concentration in the coal pillars, the coal body interior at working face and rock at roadway roof deep position” and “broken rock at roadway roof shallow position”. The technical principles of this method are as follows.

  1. (1)

    According to the detection results of the loose failure range of the roadway rock mass, the shallow broken rock mass of the roof is reinforced with high pre-stress grouting cables to improve its self-supporting capacity.

  2. (2)

    Roof directional pre-splitting technology is developed by using the rock mass characteristics of high compressive strength and low tensile strength, which is used to cut the roadway roof to relieve the pressure of the deep rock mass, and to cut off the stress transfer between the gob and the roadway roof. Broken and expansive rocks that are cut off automatically fall into the gob, forming the roadway.

  3. (3)

    Drilling pressure relief technology is used to relieve coal pressure and reduce stress concentration in the solid coal side of the roadway.

Fig. 11
figure 11

Schematic diagram of the integrated control method of comprehensive pressure relief for coal and rock masses and anchor grouting reinforcement in deep mining roadway

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To sum up, on the one hand, this method can realize the effective reinforcement of shallow rock mass and the comprehensive pressure relief control of deep coal and rock mass. On the other hand, it can automatically form a stable roadway structure and remove the retention of coal pillars.

4.2 Method validation

To verify the effectiveness of the new method, we carried out a large-scale geomechanical model based on the field geological field conditions and the pressure relief parameters optimized using numerical simulation.

  1. (1)

    Model test scheme

The size of the physical model testing system is 2400 mm × 2400 mm × 500 mm (width × height × thickness). The system is used for comparative testing of traditional reinforcement pressure relief method with and the CPR-AG method. Combined with system parameters, field formation conditions and similarity theory (Wang et al. 2022), the geometric similarity scale \({\text{C}}_{{\text{l}}}\) is \({1:30}\), the unit weight similar scale \({\text{C}}_{\upgamma }\) is \({1:1}{\text{.5}}\), and the stress similarity scale \({\text{C}}_{\upsigma }\) is \({1:45}\). Based on the similar scale and field conditions, the working face and roadway parameters in the model body are determined, as shown in Fig. 12a.

Fig. 12
figure 12

Geomechanical model testing system and monitoring

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The monitoring section is arranged in the middle of the model body, with stress monitoring points placed, as shown in Fig. 12b to continuously monitor the stress at the working face and roadway. The hydraulic loading device is used to apply a load to the model surface, bringing the horizontal stress near the working face to 0.52 MPa and the vertical stress to 0.43 MPa. The order in which the model body is mined is as follows. Step 1: Excavate the right roadway. Step 2: Excavate the working face and retain the right-side roadway. Step 3: Excavate the left roadway under the condition of coal pillar roadway protection. See Fig. 12b for details.

  1. (2)

    Testing process and result analysis

We conducted model tests in accordance with the above mining sequence and monitoring scheme, and then analyzed the stress distribution at the working face and roadway surrounding rocks using the traditional method and the CPR-AG method after Step 3, and the results shown in Fig. 13. The monitoring points in Fig. 13a are “H4, G5, F3, E1, D1, C1, B5, and A4” from left to right.

Fig. 13
figure 13

Stress distribution of the working face and roadway in the model test

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According to the analysis in Fig. 13:

  1. (a)

    Surrounding rock stress at the working face: The surrounding rock stress at the working face on the side of gob-side entry driving with coal pillars is significantly higher than that using the method of CPR-AG, as shown in Fig. 13a. The average stress of surrounding rock for the former is 0.38 MPa, and that for the latter is 0.32 MPa, which is 15.8% reduced. It shows that the method of CPR-AG effectively releases the surrounding rock stress at the working face and reduces the potential safety hazards caused by the stress concentration of the overburden.

  2. (b)

    Roadway roof stress: The roadway roof stress using the CPR-AG method is generally lower than that of gob-side entry driving method with coal pillars, as shown in Fig. 13b. The stress difference between these two methods at monitoring points are 63.0%, 47.4%, 41.1%, 36.9%, and 30.3%, respectively, with an average of 43.7%. The results show that CPR-AG method has obvious pressure relief effect on the deep and shallow surrounding rocks of the roadway roof, which is beneficial to the overall control of the roadway.

  3. (c)

    Roadway shoulder stress: The shoulder stresses when using the gob-side entry driving method and the CPR-AG method are increased first and then decreased to form a certain range of stress concentration area, as shown in Fig. 13b. The former has a maximum stress of 1.04 MPa and appears at the H2 monitoring point, 100 mm from the roadway surface. The latter has a maximum stress of 0.91 MPa and appears at the A3 monitoring point, 150 mm from the roadway surface. The latter is 12.5% lower than the former, and is farther from the roadway surface. The results show that the CPR-AG method reduces the rock stress at the roadway shoulder and transfers it to the deep stratum, which is beneficial to the rock stability in the roadway. At the same time, the peak shoulder stresses of these two methods in the numerical test are 59.40 MPa and 48.96 MPa, respectively. Compared with the physical model test data in the same position, the differences are 21.2% and 16.4%, respectively, which verifies the effectiveness of the numerical model.

5 Engineering application

The rock stratum type and geological conditions at the working face 2215 are basically the same as that at the working face 2223 in Suncun Coal Mine, both of which are deep and high stressed. The CPR-AG method proposed in this work is applied to the haulage roadway of the working face 2215. Based on the above testing results, the specific parameters were designed as follows. The roadway roof was reinforced by grouting cables (Φ 22 mm × 6000 mm), with a spacing of 1000 mm × 800 mm. The grouting materials used for grouting cable on site is cement slurry. The Portland cement with strength grade of 42.5 is selected. Water-cement ratio is 0.6, and ACZ-II type cement additive is added. The additive is 8% of the cement quality. The roof cutting height was 8 m and the roof cutting angle was 20°. The drilling length was 20 m, drilling diameter was 150 mm, and drilling spacing was 5 m. Two rows of single hydraulic props are installed in the roadway along with coal mining, which initial support force is 90 kN. Figure 14 shows the field situation after this method was applied. We chose to monitor the displacement of the roadway roof and the force of grouting cables in a typical monitoring section, as shown in Fig. 15. The negative value in the x-axis represents the distance ahead of the working face, while the positive value represents the distance behind the working face, which is the stage of the automatically formed roadway.

Fig. 14
figure 14

Field application in 2215 haulage roadway

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Fig. 15
figure 15

Evolution of roof deformation and cable force

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According to Figs. 14 and 15, we find that within − 50 m to − 10 m from the working face, both roof deformation and grouting cable force show a rapid increasing trend. Within − 10 m to 10 m from the working face, the grouting cable force drops significantly, while the roof deformation begins to get stabilized. It shows that the surrounding rock stress can be effectively released by roof cutting of the rock mass and drilling pressure relief of the coal body. The maximum deformation of the roadway roof is 154 mm in the stage of automatically formed roadway, which is 535 mm less than that at the working face 2223 mentioned above, meeting the requirements of field safety and roadway reuse.

6 Conclusions

  1. (1)

    To address the surrounding rock control problems in high-stressed deep mining roadways, we used Suncun Coal Mine with the maximum mining depth in China as the engineering background to carry out field monitoring of the traditional method of gob-side entry driving. By analyzing surrounding rock deformation, loosening range, and reinforcement system stress, we find that the reason why the surrounding rock is difficult to control in this type of roadway is due to the stress concentration in coal and deep rock mass and the low bearing capacity of shallow rock mass. Moreover, this method also has the issues of large roadway excavation and coal pillars cannot be recovered.

  2. (2)

    Combined with the field test of anchor grouting parameters, we carried out numerical simulation with multi factors on comprehensive pressure relief for coal and rock mass and anchor grouting reinforcement. We established the quantitative evaluation indexes for surrounding rock stress and displacement, and analyzed the evolution of rock stress and displacement under different pressure relief parameters (roof cutting for rock and drilling pressure relief for coal). Furthermore, we clarified the coupling control mechanism of “roof cutting pressure relief for deep roof rock + drilling pressure relief for coal body + anchor grouting reinforcement for shallow roof rock”. In addition, we obtained the design basis of coal and rock mass pressure relief parameters and anchor grouting reinforcement parameters in deep mining roadways.

  3. (3)

    We proposed an integrated control method for the comprehensive pressure relief for coal and rock mass and anchor grouting reinforcement in deep mining roadway. We used geomechanical model test to compare the traditional gob-side entry driving method and this new method, and obtained the evolution of stress at the working face and in the roadway surrounding rock. The new method can reduce the average stress at the working face by 15.8%, which verifies its advantages in the overall pressure relief control for coal and rock mass in deep mining roadways. The field design and application of the CPR-AG method are carried out, which reduces the deformation of the surrounding rock by 535 mm. This method achieves the safety control of deep mining roadways with broken roof in high-stress environment, while reducing roadway excavation and eliminating coal pillar retention.