1 Introduction

China's hard roof conditions are highly variable, with thicknesses ranging from tens to hundreds of meters, and is a relatively difficult type of roof to control, with coal resource reserves under hard roof conditions accounting for about one-third of the total reserves. At present, nearly 40% of the comprehensive mining face belongs to the hard roof working face, and more than 50% of the mines in the country have hard roof working face mining problems. There is an urgent need to solve the problem of strong mine pressure power disasters under hard roof conditions to ensure the safe and efficient recovery of working faces (Han et al. 2015; Lin et al. 2016; Pang and Zhang 2013; Tan et al. 2015).

The most common use for coal recovery under hard roof conditions in China is deep-hole pre-cracking blasting. With the wide application of blasting technology, many experts at home and abroad have done more in-depth research on deep-hole pre-cracking blasting technology, and have more or less achieved certain research results (Özel and Ünal 1998; Talhi et al. 2004; Roy et al. 2003). Xie et al. (2011) used a combination of finite element and discrete element methods to investigate the effect of nodal geometric features on pre-fracture blasting. The relationship between the pre-fracture effect and the nodal angle and nodal distance was obtained. Pomasoncco-Najarro et al. (2022) performing the simulation with the software determines a difference between conventional blasting with the design of the pre-cut technique, optimizing the parameters of the perforation mesh, selection of the explosive charge, the order of detonation. The study found a 60%–70% reduction in excess excavation, showing a better finish of the tunnel contour, yielding a 16% reduction in support costs. Kumar (2022) among the various optimization methods, the firefly algorithm proved to be a very good optimizer. Therefore, it was used to optimize the field parameters and implemented. The resulting optimized parameters showed a significant reduction in the ground vibration of 14.58%. Feng et al. (2022) used the continuous-discontinuous element method to evaluate the blasting effectiveness of a two-layer bundle hole cutting blast design. Both straight and inclined holes were considered. The optimal design was determined after comparing the damaged area and the size of the rock masses being blasted. Xie et al. (2017) applied controlled pre-cracking blasting technology to coal mine engineering and improved the permeability and gas extraction efficiency of coal seams. Li et al. (2016) analyzed the effect of blasting on slope stability by using the pseudo-static method based on the theory of limit equilibrium method (LEM) and obtained the formula for calculating the blast hole spacing of pre-cracking blasting. Based on the analysis of the principle of vibration tester and the detection of shock wave vibration by the vibration sensor, the detection scheme of blasting vibration was derived by blasting the pre-cracked and non-pre-cracked surfaces of the test object. Aliabadian et al. (2014) used the two-dimensional different cell method to simulate the pre-cracking of rock slopes (open pit mining), where the history of the blast load acts as a function of time on the inner wall of each borehole, and the important parameters considered in the analysis are the stress tensor and the fracturing pattern.

Scholars at home and abroad have carried out a lot of research on hard roof and pre-cracking blasting, and have also achieved rich results. Hard roof plate coal seam mining process to form a large span exposed structure, its bearing capacity and not easy to fall. In the load-bearing state, it accumulates high energy of elastic potential and releases energy in the form of vibration at the moment of breakage and instability, which forms an obvious instability dynamic force. This is a strong mining pressure effect on the coal and rock body around the quarry and the working face support equipment, and even induces mining pressure disaster. However, under different conditions such as geometrical parameters of the working face of the hard roof plate and the mining frontier, it makes a big difference in the appearance of the mining pressure. The main factors that influence the dynamic evolution of the space layer structure of the overburden and the key points in space and time are not clear. This paper takes the 11,129 working face of Zhangji mine as the research background. The roof plate of this working face is a 16 m thick sandstone composite roof plate, and its compressive strength reaches 89.8 MPa, which makes the roof plate strong and difficult to collapse. The increase in the roof cantilever length leads to the energy accumulation, and it is easily to cause dynamic pressure impact on the working face when it reaches the critical value and suddenly falls. It has seriously influenced and restricted the safe and efficient mining and development of high quality of coal mines, and it is necessary to artificially release the thick and hard roof plate above the coal seam to guarantee the safe and efficient production of coal mines. Therefore, for the 11,129 working face thick hard sandstone roof plate direct cover and mining technology conditions, combined with sandstone roof plate mechanical properties test results. It mainly carries out the feasibility study of deep-hole pre-cracking blasting technology for sandstone roof plate in the comprehensive mining face. The research method combining theory, physical experiment, numerical simulation and field test is adopted to systematically study the interaction relationship between the thick and hard sandstone roof plate directly covering and the bearing capacity of the support. Explore the macroscopic evolution law of deep hole pre-cracking blasting and the characteristics of mine pressure manifestation in the working face. It can make the difficult roof plate transformed into the roof plate that can fall, improve the hard roof falling, reduce the area of hanging roof in the mining hollow area, and weaken the strength of the roof plate coming pressure. This will reduce the impact of the pressure on the working face support and realize the safe and efficient mining of 11,129 working face.

2 Project profile analysis

The overall shape of the 11,129 working face in Zhangji Mine is a monoclinic structure, with locally developed secondary folds. The form of coal seams is generally high in the west and low in the east, and the yield of coal and rock seams in the working face changes greatly. Coal (rock) seams: inclination 55–170°, dip 2–6°, average 4°. There are 19 faults in the face, among which F1112902, F1112901, F1112928, F1112956, Fy204 are positive faults, which have a big influence on the mining of the working face, and the other faults have a drop of less than 2 m. Among them, the F217 (H = 0–18 m∠35°–50°) fault is the boundary fault of the northern part of the working face, and the actual position of the fault will influence the layout of the cutting eyes of the working face, and the F217 (H = 0 ~ 18 m∠35°–50°) fault is the boundary fault of the northern part of the working face. F217 (H = 0 ~ 18 m∠35° ~ 50°) is the northern boundary fault of the working face, the actual position of this fault will affect the arrangement of the cutting eye of the working face, and Fg5 (H = 4.5 m∠60° ~ 70°) crosses the working face, which has a big influence on the track running in the working face. 11,129 working face is the first 9-coal face of the east one mining area, and is adjacent to the 21,119 working face in the west (not yet under construction), which is bounded by the Fs331 in the north, the system lane of the east one mining area in the south, the unmining area of 9–1 coal in the east, and the line of the protection coal pillar of Zhangjizhen town in the west.. The overlying 11–2 coal has been mined back and the underlying 8 coal is unmined. In terms of water-filling factors, there is no direct hydraulic connection between the aquifer in the middle part of the freshwater boundary and the coal strata. The sandstone fissure water in the top plate of 9 coal is mainly static storage and has unevenness, and its water enrichment depends on the development degree of the fissures in the top plate of 9 coal seam and the interconnectivity of the fissures. The working face is mainly affected by the sandstone fissure water of the roof plate in normal excavation, and the normal water influx is expected to be about 3–5 m3/h, and the maximum water influx is 20 m3/h.

Zhangji coal mine 11,129 working face inclined length 240 m, buried depth 700 m from the surface, main mining 9–1 coal thickness 1.0 ~ 2.9 m, average thickness 1.9 m. belong to the more stable and mineable thin coal seam, coal rock components are mainly dark coal. It contains a small amount of bright coal and mirrors coal strips, and the macroscopic coal rock type is faint to semi-dark, with 0–2 layers of mudstone or carbonaceous mudstone intercalation. The overall form of the main mining coal seam is a monoclinic structure, generally high in the west and low in the east, and the dip angle of the coal seam is 2° ~ 6°. The direct top lithology is 7 m thick quartz sandstone, the old top lithology is 8.0 m thick siltstone, the direct bottom lithology is 2.5 m thick mudstone, and the old bottom lithology is 5.2 m thick siltstone. The rock assignment of the roof and floor of the coal seam is shown in Fig. 1.

Fig. 1
figure 1

Mining and excavation layout of working face

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With the extension of coal resources mining to the deep part, it faces the influence of "three highs and one disturbance". The existence of thick and hard top plate makes the 11,129 working face mining process has multiple geological pressure problems. Firstly, the thick and hard roof plate is hard and has high bearing strength, which makes it difficult for the roof plate to fall in the early stage of mining. This is one of the main problems facing the mine at present, and the gradual increase in the space of the mining airspace easily leads to the impact ground pressure. Secondly, the sudden fall of the roof plate in the process of cyclic mining makes the working face and the two roadways appear dramatic mineral pressure, leading to the risk of coal wall flaking, roof fall and pressure frame. Therefore, for these geological features, it is necessary to take targeted measures to make sure the mine production is safe and efficient.

3 Theoretical analysis of the structural motion of the overburden slab in the mining space

3.1 Mechanical analysis of top slab breakage deflection at the beginning of mining

Li et al. (2008) regarded the hard roof of the mining ground as an elastic thin plate, determined the theoretical calculation formula of the initial and periodic collapse step of the roof using the displacement variation method, as well as applied it to engineering practice. Jia et al. (1999) established the corresponding elastic thin plate mechanical model according to different support boundary conditions to calculate the predicted incoming pressure step and incoming pressure strength of the top plate. The results of previous studies show that the thin plate theory is more applicable to solving the problem of rock control on the roof of the quarry. In this paper, we use theoretical analysis to quantitatively analyze the law of direct top initial breakage and periodic breakage, so that we can analyze the change of deflection during the collapse of the top plate.

Mining thin plate theory is to treat the direct top rock layer before fracture as an elastic thin plate structure, with the mid-plane of equally divided thickness h as the xy-plane and vertically downward as the z-plane. When the thin plate is bent, the deflection deformation occurs at each point in the direction perpendicular to the middle plane, and its deflection deformation ω(x,y) is much smaller than its thickness h. In line with the basic assumption that the displacement and deformation are tiny, it belongs to the small deflection bending theory, and the established mechanical model is shown in Fig. 2 below.

Fig. 2
figure 2

Mechanical model of initial mining breakage of rock formation

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Small deflection differential equation for elastic thin plate (Li et al. 2008):

$$D\nabla^{4} \omega = q$$
(1)

Which

$$D = \frac{{Eh^{3} }}{{12\left( {1 - \mu^{2} } \right)}}$$
(2)

where:

D—The bending stiffness of the thin plate. \(\nabla\)—Laplace operator.

q—Load in each unit area of the thin plate. E—modulus of elasticity of thin plate.

h—general thickness of the thin plate. μ—Poisson's ratio.

Combined with the mechanical model, it is considered that the top plate at the beginning of mining can be regarded as a four-sided solidly supported thin plate structure. The initial breakage deflection of the elastic thin plate ω:

$$\omega = \frac{{qa^{4} \left( {1 - \cos \frac{2\pi x}{a}} \right)\left( {1 - \cos \frac{2\pi y}{b}} \right)}}{{4D\pi^{4} \left[ {3 + 3\left( \frac{a}{b} \right)^{4} + 2\left( \frac{a}{b} \right)^{2} } \right]}}$$
(3)

According to the mine pressure law and 11,129 working face geological structure conditions to obtain engineering parameters, direct top thickness: h = 7.1 + 0.5 = 7.6 m, face length b = 240 m, μ take 0.3, top plate elastic modulus E = 33.4GPa. After pre-cracking and blasting, the top plate of the working face is broken for the first time for 45 m, with a cycle break step of 12.5–13.5 m, averaging 13 m, and a burial depth of H = 700 m. Excluding external and human factors, different parameters are taken to compare and analyze the relationship between the deflection changes of the top plate under different influencing factors.

The values of Table 1 below are substituted into Eq. (3) for the solution. Compare and analyze the deflection variation of the initial breakage of the top slab under different influencing factors.

Table 1 Table of values of initial mining breakage factors
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As shown in Fig. 3, a small space is formed under the roof plate at the initial mining stage of the coal seam, and the board structure of the roof plate shows a four-side fixed-supported plate structure, the roof carries the weight of the overlying coal and rock layers, and the bottom is not limited. With the advance of the coal seam, the roof slab and its overburden rock layer show a downward concave tendency, which is shown that the deformation degree of the middle part of the roof slab is larger than that of the four sides of the roof slab. Different physical parameters of the roof slab and the distance of coal seam advancement also have different effects on the degree of roof deformation.

Fig. 3
figure 3

Deflection curve of top plate with different factors

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In the mining area, the length of the working face b is usually 200–300 m, and the change of the face length is positively correlated with the bending deformation of the roof plate. When the working face advances 45 m and the face length b = 250 m, the deflection in the middle of the working face is 393 mm, and the deformation gradually decreases when it is close to the surrounding coal wall, and finally the roof plate acted on the coal seam. The influence of high ground stress is especially obvious with the increase of the mining depth of the coal seam, and the high stress and mining disturbance increase the chance of breaking and falling of the roof slab and the suddenly releasing of the elastic energy. The shallow endowed coal seams (within 300 m depth) are relatively weakly affected by ground stress, and the deflection of the middle part of the roof slab is only 23.3 mm. When the mining depth gradually increases from 300 to 900 m, the deflection in the middle of the roof increases to 227.7 mm. It can be seen that the conditions of the coal seam itself greatly restrict the mining of coal, and when the depth increases by 600 m, the degree of roof deflection increases by about 10 times. The physical properties of the roof plate (modulus of elasticity and thickness) are also important factors affecting the bearing capacity of the roof plate. Thickness of the roof plate is large, high strength, difficult to fall, and the relative deflection deformation is small. The mechanical parameters of the roof plate should be measured in advance before mining the working face.

3.2 Mechanical analysis of the period break deflection of the roof plate

The cycle breaking model is shown in Fig. 4. Similarly, the periodic breakage can be viewed as a thin plate model with solid support on one side and free on the other. The equation of the periodic breaking deflection curve (Li et al. 2008):

$$\omega = \frac{{qab\left( {\pi - 2} \right)\left( {1 - \cos \frac{\pi x}{{2a}}} \right)\left( {1 - \cos \frac{2\pi y}{b}} \right)}}{{\pi \left[ {D\pi^{4} \left[ {\frac{3b}{{64a^{3} }} + \frac{\pi - 4}{{2\pi ab}} + \frac{{4a\left( {3\pi - 8} \right)}}{{\pi b^{3} }}} \right] + \frac{{D\pi^{3} \left( {\pi + 4} \right)\left( {1 - \mu } \right)}}{2ab}} \right]}}$$
(4)
Fig. 4
figure 4

Rock cycle breakage mechanical model

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The different values of Table 2 below are substituted into Eq. (4) for the solution. Compare and analyze the deflection variation of the top plate cycle breaking under different influencing factors.

Table 2 Table of values of cycle breakage factors
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As shown in Fig. 5, as the coal seam advances further, the roof plate is cyclically falling, and the roof plate is changing from a four-side fixed support to a three-side fixed support with one side free. At this moment, the maximum value of roof deflection is moved from the center to the free end, which is shown as the roof plate rotating and falling down to the free end. The influence of underground stress and rock parameters on the degree of roof deformation in the process of cyclic mining is basically the same as the initial falling rule. When the cycle breaking step is 13 m and the face length b = 250 m, the deflection of the free end of the roof plate is 56.7 mm.Different breaking lengths of the roof board have more obvious influence on its deflection, when the cycle breaking a = 16 m, the deflection of the free end of the roof board is 125.9 mm, which can provide the basis for the design of the spacing of the field holes. The physical properties of the rock layer (modulus of elasticity and thickness) are also important factors affecting the difficulty of rock collapse. The characteristics of the rock layer are the result of the surrounding environment, and its deformation rule is similar to the change rule of the roof deflection during the initial mining period.

Fig. 5
figure 5

Deflection curve of roof plate with different factors

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Comprehensive analysis of the above, the bearing capacity of the working face roof and its degree of deformation are mainly affected by the ground stress (depth of the coal seam) and the physical properties of the rock layer itself (modulus of elasticity and thickness). The less important reason is the length of the working face arrangement and the influence of its disturbance in the process of coal seam mining. The mechanical parameters of the roof rock layer can be measured by drilling and coring before mining the coal seam, and then the bearing ability of the roof rock layer can be inferred. It can provide theoretical basis support for the following precracking and blasting roof control and mining support selection.

3.3 Mechanical analysis of the relationship between roof deflection and support load-bearing

The change of brace resistance is a transient indicator to describe the working condition of brace. The brace stiffness is a permanent indicator to analyze the overall stability of the system in the process of "support -rock" interaction, so it is more important to analyze the "support-rock" relationship. The sinkage of the top plate is mainly determined by the top plate load and the stiffness of the system consisting of " roof-support-floor", and the system stiffness is calculated as (Xu 2015):

$$\frac{1}{K} = \frac{1}{{K_{f} }} + \frac{1}{{K_{z} }} + \frac{1}{{K_{d} }}$$
(5)

where, K is the system stiffness, MPa/m; Kz is the support stiffness, MPa/m; Kf is the top plate stiffness, MPa/m; Kd is the roof stiffness, MPa/m.

The modulus of elasticity and the stiffness before the damage of common rocks in the roof and floor of the recovery workings are given in the literature in Table 3. The general rock modulus of elasticity is 8 to 40 GPa, and the stiffness of the top and bottom plates before damage is 8–4 GPa/m. The current range of bracket stiffness is about 10–80 MPa/m, and the roof and floor stiffness is 100–4,000 times the bracket stiffness. From Eq. (5), the stiffness of the whole system is mainly determined by the support stiffness, while the roof and floor have less influence on the system stiffness.

Table 3 Common coal rocks' stiffness (Xu 2015)
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For the long wall working face of initial mining and cycle advance, if the direct top is in front of the coal wall and the roof control area is without fracture, an infinitely long elastic base continuous rock slab is formed above the coal body, and the support. The direct top in front of the coal wall is a semi-infinite continuous plate with solid coal as the elastic base. The direct top behind the coal wall is for a semi-infinite continuous plate with the support as the elastic base. Combined with the characteristics of initial mining and periodic breakage, the mechanical relationship of "roof-support-floor" is analyzed, as shown in Fig. 6.

Fig. 6
figure 6

Schematic diagram of dynamic load response of working face support

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In the process of advancing the coal mining face, the roof transport and bracing forces can be divided into the following stages:

  1. (1)

    Initial stage of mining This is the stage when the working face has just started to advance. In this stage, the rock layer above the roof is affected by the disturbance of coal mining work. Small local fissures and displacements will appear, but they will not cause the breaking and rotation of the rock layer. The bracing support effect is small.

  2. (2)

    Rotating and sinking stage When the working face advances a certain distance, the rock seam above the coal seam starts to be stripped and forms the off-set. In this stage, by the action of geological stress, the rock seam fissures develop and penetrate, and start to produce overall sinking and tilting. At the same time, the stress concentration zone of the top slab is gradually formed by the compaction effect of the stripped rock layer. The brace bears the pressure from the sinking of the top slab and provides a certain support force.

  3. (3)

    Breakdown and fragmentation stage When the working face continues to advance, the fracture zone formed gradually develops and expands, the overhanging roof reaches a certain length, and the roof plate rotates and sinks substantially. In this stage, the roof plate changes from continuous to localized, and the bracket is under pressure from the rotating fracture of the roof plate and the coal and rock body above it, which needs to provide enough support to stabilize the roof plate.

In order to grasp the interaction of the "roof-support-floor" system more accurately from the theoretical aspect. By calculating the bearing capacity of the bracket in the roof control area, the degree of rotation and sinking of the roof plate can be determined. At present, the range of bracket line stiffness (Xu 2015) is 100 ~ 600 kN/m, and the range of stiffness is 10 ~ 80 MPa/m. The analytical formula of bracket-bearing capacity is:

$$F = KA\omega$$
(6)

where: K—bracket stiffness, MPa/m; A—support surface area of the bracket, m2.

ZZ10800/18/38D type support cover type bracket is used in the working face of the quarry, and the rigidity of the bracket is 25.0 MPa/m. The top control distance of the support is 5275 ~ 6175 mm, which is 5.7 m. The width of the support is 1.5 m, so the top control area A = 5.7 × 1.5 = 8.55 m2. Substitute Eq. (3) and Table 1 into Eq. (6) to solve for the bearing characteristics within the roof control interval of the bracket under different factors during the initial mining period, as shown in Fig. 7 below.

Fig. 7
figure 7

Curve of the influence of different factors on the working pressure of the support during the initial mining

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The extent to which different geologic conditions cause the roof to settle and break is an important consideration that indirectly affects the amount of pressure supported by the mine's supports. The coal mining gradually excavates deeper and the increase of the surrounding rock stress is the root cause of the intensified movement of the rock seam. With the increase of coal mining depth, the pressure at the minimum controlled roof distance gradually increases from 11.5 to 34.6 MPa. The thickness and hardness of the roof rock can reflect the bearing capacity of the overlying coal seam. Under the condition of a certain thickness, the increase of rock elastic modulus (hardness) increases the bearing capacity of the rock stratum, and the force acting on the support is relatively reduced. The modulus of elasticity increases from 8 to 32GPa, and the force on the bracket is reduced by 75%. In the same way, the rock thickness and the bearing capacity of the support change positively under the same hardness. The breaking step of the roof also depends on the characteristics of the lithology to some extent, and the influence of the length of the working face on the collapse of the roof is relatively small. However, there is a small increase in the growth rate in the roof control area (Fig. 8).

Fig. 8
figure 8

Curve of the influence of different factors on the working pressure of the support in cycle recovery

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After experiencing the initial collapse the roof plate becomes free at one end from four solid sides. Similarly, formula (4) and Table 2 are substituted into formula (6) to solve the bearing characteristics within the bracket-controlled roof interval under different factors during the cycle breakup. The law of the cycle collapse stage is basically the same as the initial one, when the strength of the roof plate is increased from 8 to 32 GPa, the bearing capacity of the roof plate on the overlying coal rock body increases, and the stress acting on the quarry support decreases from 54.1 to 13.5 MPa, which is about three times less than the initial stage. When the modulus of elasticity of the roof plate is 8 GPa, the minimum and maximum roof control distance pressure of the support is 54.1 MPa and 81.9 MPa, with a fluctuation degree of nearly 28 MPa. With the increase of roof bearing capacity to 32GPa, the minimum and maximum roof control distance pressures are 13.5 MPa and 20.5 MPa, the fluctuation degree is reduced to 7 MPa. The high stress in the deep part and the physical properties of the rock layer itself (hardness and thickness) are the main controlling factors to determine the bearing capacity of the support. Artificially controlling the length of the coal mining face indirectly regulates the bearing capacity of the support to a certain level, but it is relatively small compared with the influence of other factors.

4 Design of strong mining to pressure cut roof control scheme for hard roof

4.1 Analysis of roof rock loosening blasting parameters

The mining stress and rock movement will cause local stress concentration, change the stress state of the mining envelope, the distribution and properties of fractures in the mining envelope and overlying rocks, etc. It provides the conditions and triggering factors for coal and gas protrusion and impact pressure to occur. In the prediction and evaluation of coal and gas protrusion and impact pressure, etc. The first thing is to study the distribution and evolution law of mining stress in the area to be evaluated, and the law of rock fracture and movement. Especially, the hard roof breakage and movement law (Wang 2019). Therefore, it is necessary to rationally and scientifically determine the "three zones" at the stage before coal seam mining.

According to the theoretical results of previous studies and mine log data. The height of the collapse zone and fracture zone under hard roof conditions can be determined by the following equation (Chang 2015):

$$\left. \begin{gathered} H_{m} = \frac{100M}{{2.1M + 16}} + 2.5 \hfill \\ H_{v} = 30\sqrt M + 10 \hfill \\ \end{gathered} \right\}$$
(7)

where: M is the mining height of the working face, m. Hm is the height of rock fracture development above the mining area, m. Hv is the height of rock fracture zone yurt above the mining area, m.

At home and abroad, blasting forced roof release technology mainly has the following two kinds: one is blasting forced roof release along the roof of the cuttings and along the wind tunnel, and the other is over-advance deep hole blasting roof release in the machine tunnel drilling parallel working face. The over-advance loosening method is highly adaptable to the geological and technical conditions of the coal seam. It has less influence on the normal recovery of working face, requires less special equipment, and makes the roof plate with good integrity change to the roof plate with artificial fissure development. Make the top plate which is not easy to fall change to the top plate which is easy to fall, reduce the overhanging roof area in the mining area, and eliminate the pressure gathering. Forced roof release makes the formed fissure zone must have a certain width, the width is large, and the bubble fall is obvious. But the width is too large, which may cause the roof plate to break. The roof plate is difficult to manage during coal mining, and the workload of drilling and blasting increases exponentially. In order to ensure that the hard roof can fall in time, it is necessary to determine a reasonable blasting width, the number of rows of shell holes. The blasting width is related to the length of the cuttings. When the length of the eye is short, the top plate is subject to strong four-sided clamping, the stability of the overhanging top plate is better, and the blasting fissure zone is relatively large. After blasting, the fissures intersect each other in circles, forming a web of fissures.

There are many formulas for calculating the radius of blasting fracture circle. One of the most commonly used formulas is (Dai 2001):

$$R_{p} = \left( {\frac{{\nu PS_{t} }}{1 - \nu }} \right)^{1/\alpha } r_{b}$$
(8)

Which

$$P = \frac{1}{8}\rho_{0} D^{2} \left( {r_{c} /r_{b} } \right)^{6} n,\;\;\alpha = 2 - \nu /(1 - \nu )$$

where: Rp is the radius of the loosening circle. P is the initial radial stress peak of the stress wave. α is the stress wave decay value. d is the explosive burst velocity. ρ0 is the explosive density. rc is the radius of the package. rb is the radius of the borehole. St is the tensile strength of the rock. v is the Poisson's ratio. n is the pressure increase factor, 8–11.

4.2 Design of roof cutting and drilling scheme

The 11,129 working face uses two-way poly energy blasting pre-cracking technology. Specific sizes of explosives are loaded in two set directions in the energy gathering device with the effect of energy gathering. After the explosives detonation, the shell hole surrounding rock in the non-set direction of uniform pressure, while in the set upward concentration of tension, relying on the characteristics of the rock resistance to pressure and fear of tension, so as to realize the blasting rock block according to the set direction of tension fracture molding.

Two-way energy blasting pre-cracking technology is based on a variety of energy blasting and directional blasting methods developed on the basis of a new type of energy blasting technology. Construction process is simple, the application only needs to be constructed in the pre-fracture line of the holes, the use of two-way poly energy device charge, and make the direction of poly energy corresponds to the pre-fracture direction of the rock body. The blast product will form a poly energy flow in two directions and produce concentrated tensile stress, so that the pre-cracked holes run along the direction of the energy gathering, forming a pre-cracked surface. As the rock between the drill holes is broken, blasting explosives unit consumption will be greatly reduced, and at the same time, because of the protection of the surrounding rock by the poly energy device, the damage to the rock body surrounding the drill holes is also greatly reduced, which can achieve the realization of pre-cracking and at the same time protect the roof of the roadway.

According to the specific geological situation of 11,129 working face, it was decided to adopt a segmental deep hole pre-cracking blasting program for the working face roof. Due to the thick hard sandstone straight overlying roof plate and composite roof plate in the two down-slots at different distances from the cutting eye. Therefore, in order to effectively control the influence of the roof slab on the mining pressure manifestation in the working face mining process. It is proposed to design different deep hole loosening blasting programs for the working face roof plate in sections. The 200 m position of the retreat point was selected as the dividing line to design different deep hole blasting schemes for the roof plate. The flat section of the shell hole is shown in Fig. 9 blasting (Tables 4, 5 and 6), breaking program is divided into three steps:

Fig. 9
figure 9

Working face gunhole arrangement diagram

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Table 4 Table of blasting parameters at the working face cut-hole
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Table 5 Table of blasting parameters in the range of 0 ~ 200 m at the working face
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Table 6 Table of blasting parameters in the range of 200–700 m on the working face
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The first step is to cut the intraocular blast and shorten the initial incoming pressure step.

The second step was to blast the roof above the interior of the workings within 200 m. Solve the problem of the old roof giant thick sandstone above the composite roof slab.

The third step is a blasting program to the roof above the interior of the workings within 200 to 700 m. Solve the giant thick sandstone direct overburden problem.Deep hole blasting is carried out in the cutting hole to force the roof release before the installation of the support at the working face. The advantages of this plan are that it does not affect production, can effectively cut off the roof and shorten the initial incoming pressure step. In the cut hole, holes are drilled 35 m away from the transport smooth and 1.5 m away from the coal gang on the mining side. Among them, 1-1#, 1-2#, 1-3#, 1-4#, 1-5# holes are spaced 20 m apart. The holes are punched at 40 m from the track and 1.5 m from the coal gang on the mining side. Among the holes, 2-5#, 2-4#, 2-3#, 2-2#, 2-1# holes are spaced 20 m apart.

Within 200 m in front of the working face, as shown in Fig. 9. The first group of shell holes is 12 m away from the coal wall of the cuttings, with a horizontal spacing of 1.5 m and staggered long and short holes, with a group spacing of 20 m (the first shell hole of the second group is 20 m away from the third shell hole of the first group). Deep-hole loosening blasting was carried out in the track chute and transport chute to the roof plate above the inner part of the working face. According to the characteristics of roof collapse and mine pressure at the working face, determine whether it is within the range of 100 m from the coal wall of the cuttings. Deep hole blasting is carried out to the roof plate above the inner workings. There is a high extraction lane about 30 m above the top plate of the coal seam about 30 m from the track, to avoid affecting the gas extraction.

In the range of 200 ~ 700 m in front of the cutting face of the working face. Deep hole loosening blasting was carried out every 30 m to the top plate above the inner part of the working face in both chutes. The horizontal spacing of shell holes in the group is 1.5 m, and the long and short holes are staggered. There is a high extraction lane arranged about 30 m above the top plate of the coal seam about 30 m along the track to avoid affecting the gas extraction.

By the combination of blasting on the roof plate of the cutter hole, overhead deep hole blasting in the two lanes and overhead deep hole blasting in the face to cut the roof. Make the roof plate with good integrity change to artificial fissure roof plate. Make the artificially intervene the top plate of huge sandstone which is not easy to fall and then fall section by section. Reduce the overhanging roof area in the mining area, eliminate pressure gathering, reduce the risk of roof fall and gas accumulation during recovery.

In the selection of mining explosives, water-gel explosives a new type of explosives, because it has a series of advantages such as good water resistance, less smoke after the explosion, it is suitable for high gas mining, low gas mining, high gas areas and blasting operations with coal (rock) and gas protrusion hazardous working face. According to the relevant provisions of WJ/T 9066-2018 "Coal Mine Permitted Gas Extraction Hydrocolloid Pillar Explosives", coal mine permitted three-stage emulsion explosives are used, and the parameters of explosives determination are shown in Table 7.

Table 7 Parameters of explosives
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Notes on the transportation, storage and use of explosives:

  1. (1)

    The product should be carried out in transportation, storage and use in accordance with GB6722 "Blasting Safety Regulations" and other relevant regulations.

  2. (2)

    The product is packaged in plastic film cartridges, so there is no aggregation effect. Therefore, when loading in the eye of the gun, we should pay attention to clean coal dust, rock chips and other debris to ensure that the good contact between the drug rolls to avoid refusing to explode.

  3. (3)

    the product should be used in the drug temperature above 0 ℃, so when used in cold areas, the explosives are first deposited underground for a certain period of time, when the drug temperature rises to 10 ℃ or more before use.

  4. (4)

    products in the charge length is not appropriate, you can cut the use of sharp knives in the cutting, and leave the smallest pimple knot (pimple knot large impact on the sympathetic detonation).

  5. (5)

    Found that the appearance of the product is abnormal, should stop using.

  6. (6)

    The shelf life of the product from the date of completion of the manufacture of 6 months, over the period of time the product should be fully tested the technical performance, in line with the above standards can still be used.

Deep hole blasting due to deeper holes, in order to facilitate loading, the need to use fire-retardant anti-static grooved blasting cylinders. Each section of the blasting cylinder is 2 m long, the first explosive into the blasting cylinder, and then fill the hole with explosives, in each section of explosives into the hole at the same time, the detonator will be pushed into the groove blasting cylinder, as shown in Fig. 10.

Fig. 10
figure 10

Schematic diagram of the structure of positive charge yellow mud sealing hole

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5 Characteristics of the stress field and evolution pattern of blast-controlled roof overburden

5.1 Numerical analysis of stress field before and after blasting

FLAC3D can better simulate the mechanical properties of damage or plastic flow of geologic materials when they reach the strength limit or yield limit, and it is especially suitable for analyzing progressive destructive instability and simulating large deformations. Based on the above advantages of FLAC3D in geotechnical engineering. It is not possible to monitor the movement and stress distribution characteristics of coal and rock seams in real time over a large area on site, so as to establish a numerical model to study the trend of the change of the roof stress field in the process of coal mining, and to be able to recognize the stress evolution law around the quarry from an overall perspective. It plays a certain reference value for theoretical research and field construction.

FLAC3D numerical simulation software was used to establish the numerical model. The dimensions of the model are 640 m × 650 m × 168 m. The parameters of the model and the top and bottom coal seams are shown in Fig. 11. The boundary of the model except for the top is set with fixed horizontal displacement constraints. The Mohr–Coulomb damage criterion is used for the mechanical properties of the model coal rock. Scholars at home and abroad have conducted a lot of experiments and researches on the support effect of the equivalent filled rock body in the mining area (Zhang et al. 2015, 2020; Li et al. 2019; Mahdi 2012). The mining void area was simulated using the Double-yield model. The mechanical parameters of the rock body in the Double-yield extraction zone, as shown in Table 8 below.

Fig. 11
figure 11

3D numerical simulation model

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Table 8 Mechanical parameters of the main rock materials in the mining area
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Simulation of 11,129 working face advancing along the strike (positive y-axis direction) from the open-cutting. The open-cut is 200 m away from the boundary to eliminate the influence of the boundary effect. The excavation step is 20 m and fill the gob. Advance 180 m along strike and stop mining at y = 380 m. The pressure distribution in the quarry before and after the top cutting and the change of real-time working resistance of the bracket at different positions are analyzed during the back mining process. As shown in Fig. 12.

Fig. 12
figure 12

Simulation of vertical stress change before and after roof cutting at working face

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As shown in Fig. 12. As the working face advances, the mining space is formed and the stress is redistributed. The coal rock body around the mining void area forms a support pressure zone higher than the original rock stress. The stress peak is generated in the range of 10–50 m in front of the coal wall, and the uninterrupted incoming pressure affects the operating environment of the working face. The high-energy shock wave generated by the explosive formed a blast wave along the direction of the polygonal seam, which pre-cracked the rock. The integrity of the rock formation is destroyed, forming a local block of rock. After blasting the overrun support pressure is reduced and its peak is shifted towards the deep front for the cut block. This method is good to improve the situation of overloading the support at the working face.

During the process of coal seam mining, roof blasting and roof cutting simulated in FLAC3D, the support pressure will change with the advancement of the mining face and the influence of roof blasting. The working support pressure data before and after roof blasting are extracted for comparison and analysis. The change of support pressure before roof blasting can be roughly divided into two stages: (a) the initial stage of mining. Coal seam mining at the beginning of the formation of the mining space is relatively small, the roof and bottom plate to the mining support pressure is small. (b) Mining cyclic stage. With the advancement of coal seam mining, the formation of the mining air space gradually increases, the roof-bottom plate coal rock body occurs rotary sinking action to the support, support pressure increases. When the support moves, the pressure value changes with the movement of the support, showing cyclical fluctuations. At the same time, due to the destabilization of the rock layer caused by the mining of the coal seam, the roof may undergo deformation, fracture and other kinds of activities, thus causing sudden and irregular changes in the support pressure. After the roof is blasted, the integrity of the roof is broken, and localized rock blocks are formed to act on the support, and the support pressure will rise locally.

As shown in Fig. 13, the reason that the front section of the support is closer to the coal wall, the bearing gravity is relatively small, and the support pressure is relatively small. The pressure value of the front section of the support before blasting is 9.7 MPa, and the pressure value of the front section of the support after blasting rises to 13.2 MPa. The carrying capacity of the roof is weakened after blasting, and the force acting on the support is more obvious, which is specifically reflected in the increase in the pressure of the support in the quarry. In the position of 5.35 m from the coal wall before and after blasting support pressure value is the same, 5.35 m from the coal wall, support pressure value of the rate of increase slows down, from the end of the support pressure peak value of 19 MPa, but cut the roof of the pressure peak value of 28.6 MPa, the pressure is reduced by 33.6%.In the theoretical analysis, the relationship of "roof-support" interaction is explored, and the theoretically solved support pressures at different locations in the cycle stage are highly consistent with the simulation results. The peak pressure of the theoretical calculation at the end of the support is 20.5 MPa. The peak pressure of the theoretical calculation at the end of the stent is 20.5 MPa, which is slightly higher than the simulation result within the permitted error range, and the accuracy of the simulation and the reliability of the theoretical calculation result have been verified each other.

Fig. 13
figure 13

Graph of pressure change of support before and after blasting

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5.2 Experimental study of physical similarity of overburden movement

5.2.1 Similar model experimental design

In view of the characteristics of similar material simulation, the simulated test should satisfy: geometric similarity, motion similarity, dynamic similarity, boundary condition similarity, and proportionality of corresponding physical quantities, therefore:

  1. (1)

    The model similarity ratio αl = 100 and the tolerance ratio αr = 1.67.

  2. (2)

    The model to scene time ratio αt = 1/12.

  3. (3)

    Stress similarity ratio ασ = αl × αr = 167.

The dimensions of the experimental table were L × H × W (300 cm × 130 cm × 30 cm) steel frame. The model was made with the geomechanical parameters and mining conditions of the working face of the site to be simulated as the prototype. Taking the experimental conditions into account, the self-gravity stress field of part of the topsoil layer is compensated for the uniform load.

As shown in Fig. 14 above, the stress sensor used for the test was the BX120-100AA resistance strain gauge. It was used to observe the stress changes in the coal seam and surrounding rock. The CM2B static strain gauge data acquisition system was used for dynamic acquisition and analysis. The displacement measurement points are placed on the surface of the model using the cross-pointing method to observe the changes of the overburden displacement field. The smart camera Canon EOS550D was used to take photos of the measurement points and analyze the displacement changes through photos.

Fig. 14
figure 14

Model stress test collecting system

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5.2.2 Macroscopic breaking and deformation characteristics of overburden after roof weakening

The test was carried out from the left side of the model frame at 30 m, and the excavation progressed to the right in turn. As the working face advances, the length of the overhanging roof plate in the mining area increases continuously. The load of the roof plate itself and the stress of the rock layer above are transferred downward, the rock layer produces fissures and develops. When the load-bearing limit is reached, the basic roof appears to break and sink. The macroscopic breaking characteristics are shown in Fig. 15.

Fig. 15
figure 15

Damage deformation law of quarry overburden after roof weakening

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As shown in Fig. 15. When the working face advanced to 40 m, the basic roof broke down for the first time. The collapse step is 40 m, and the collapse height is about 11 m, which is about 2.75 times of the mining height. When the working face advanced to 90 m, the roof slab was periodically collapsed and sunk. The roof slab at the middle and high level became delaminated and several longitudinal and transverse fractures appeared. The roof slab shows "oblique cut" collapse along the weakening of the roof slab. It indicates that the overlying rock layer collapsed extensively. The weight of the rock layer itself is transferred to the lower rock layer. The overburden collapse height is 44 m, about 11 times of the mining height.

When the working face was advanced to 130 m, the mining gob area was filled and compacted by broken rocks. The high rock layer falls and the rock cracks develop as high as 70 m. The interlayer cracks of the fallen rock in the mining area tend to close. The high rock cracks continue to develop. After the thick hard roof plate is weakened, the overlying rock stress of the quarry is released. The roof plate achieves the effect of pressure relief, and the rock layer falls in time. The roof plate shows cyclic falling, and the overburden collapse height is 69 m.

When the working face advanced to 160 m. Cracks appeared between the high rock layers in the mining area, with a length of 30 m and a height of 20 m. The cracks of the rock seam develop up to 45 m above the coal seam, but the cracks of the high rock seam tend to close. The trend of development to the roof seam is maintained. The stress of the upper rock seam is transferred to the key rock layer, and the thick hard roof completely collapses and sinks, the mining gob area is gradually filled by the falling gangue.

It can be seen that the overlying rock tensile stress of the mining site decreases from the lower to the higher rock level. It indicates that the surrounding rock starts to break down and deform from the low level, moving upwards to break.

5.2.3 Characteristics of rock transport and stress evolution after roof weakening

In order to further analyze the movement of mining on the surrounding rock and stress evolution characteristics. The data of one side point at 5 m above the roof plate of 9 coal is extracted, as shown in Fig. 16a. A layer of strain gauges was laid at 10 m above the 9 coal roof, and the measurement points were numbered from left to right, from 1# to 14# respectively. The strain data were exported and processed after the workface was retrieved. As shown in Fig. 16b.

Fig. 16
figure 16

The amount of rock subsidence and stress characteristic curve after the weakening of the roof slab

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From Fig. 16a, it can be seen that when the working face was retrieved less than 40 m, the overburden rock did not appear to be obviously falling. When advancing to 40 m, the roof first failed, with a maximum settlement of 3.4 m. When advancing to 90 m, 44 m of rock above the coal seam began to move and deform. The key rock layer became unstable with a maximum sinkage of 3.6 m. When it has been advanced to 160 m. The rock layer is damaged from the bottom up and sinks extensively. The degree of collapse gradually decreases from the bottom to the top. The whole form of rock movement and destruction shows symmetry.

It can be seen from Fig. 16b. When the working face advances 40 m, stress concentration occurs in front of the working face with peak stress of 22.7 MPa. The stress in the gob area is reduced. The pressure relief zone is formed within 30 m from the coal wall to the rear. The trend of stress change is basically the same for 90 m and 40 m. The high rock layer is less affected by the mining movement of the working face. The degree of damage is much smaller than that of the lower rock seam. The overrunning support pressure is 24.5 MPa respectively. When the working face was advanced to 160 m, the stress peak increased slightly. The stress variation in the extraction area is more discrete than the original rock stress.

6 Analysis of pressure release effect of deep hole blasting at the project site

6.1 Inspection of deep hole blasting effect

The rock inside the hole in a blasting operation is subjected to the shock waves and explosive gases generated by the explosion. As the blast wave propagates through the rock, it generates a large number of stress waves and shear forces within the rock, resulting in cracks and fractures within the rock. Therefore, by observing the cracking of the rock in the hole. Analysis of the crack length, number, density, morphology and orientation of the rock in the hole subjected to the blast wave. Examine and evaluate the crack extension of the rock, the state of the residue after the blast, and the presence of unstable areas and areas of risk. Helps to detect the presence of uncollapsed hazardous areas and the possibility of falling rock masses as well as to take action to prevent injury and damage (Fig. 17).

Fig. 17
figure 17

Rock crack expansion hole map

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The crack length can reflect the distance and intensity of the blast wave propagation inside the rock body. The longer the length indicates the greater the energy of the blast wave and the more severe the damage to the rock block. The number and density of cracks can reflect the distribution of cracks inside the rock mass. The higher density indicates the more serious damage of the rock body. In addition, the shape and direction of the crack can also reflect the damage mode and damage direction of the rock mass. When the blast wave propagates inside the rock mass, the stress inside the rock mass exceeds the compressive strength of the rock under the action of the blast wave. Resulting in plastic deformation and fracture of the rock, and along the direction of propagation of the blast wave shows a more obvious direction. At the mouth of the borehole, uneven rock blocks with high fragmentation appear at the borehole affected by the blast wave. Cracks in the hole at 2.3 m showed annular development, and at 3 m the cracks extended along the drill hole strike. Polyenergy cuttings played a good effect. The annular and strike cracks alternate at 4–5 m. With the depth of the probe, the crack development is mainly in the strike direction. The hole wall as a whole is relatively smooth, indicating that the overall strength of the rock in this section is high. The rock spalling in the hole at 20.8 m deep indicates that the overall strength of the rock is relatively low here.

Probe hole imaging showed that the deep hole pre-fracture blasting effect achieved better results. The relative strength differences of the rock mass at different depths were found, which provided a reference for the improvement and optimization of the blasting program.

6.2 Analysis of mine pressure analysis at the working face

In view of 11,129 working face 19–23.5 m thick sandstone roof and mining technical conditions, combined with the sandstone roof mechanical properties test results. Comprehensive analysis of the strength and stability of the giant thick sandstone roof plate. It is necessary to observe the characteristics of mine pressure during the initial release and normal recovery of the working face.

Anhui Sensing mining digital pressure gauge is used. The pressure data from the working face is collected and transmitted to the surface computer for data processing using a collector. Mine pressure observation on the working face is divided into 17 stations for analysis. Each station is equipped with 1 pressure gauge, which is connected to the front and rear pillars of the support, totaling 17 pressure gauges. Then the data is collected to the mainframe of the ground monitoring system for data management. Finally, the data is processed and displayed by the computer. The initial and periodic 3D mine pressure distribution is shown in Figs. 18 and 19.

Fig. 18
figure 18

Pressure distribution at the working face at the beginning of mining after the weakening of the roof

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Fig. 19
figure 19

Pressure distribution at the working face of the mining cycle after the weakening of the roof

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Deep hole pre-cracking blasting was carried out on the roof at the early stage of mining. At 45 m of advance, the initial pressure came on and the pressure step was determined to be 45 m. The initial pressure step of the roof plate was effectively shortened, and the peak pressure of the support was 32 MPa at this time. The pressure coming from the supports at the two ends of the working face was relatively slow. The reason is that the support is located at the end of the quarry, the bending and sinking of the top plate end is less, and the force transferred to the support is relatively small. The pressure area located in the lower and middle 60# to 70# supports is more concentrated and generally higher pressure. At the early stage of mining, according to the site observation, no abnormal conditions such as slab, roof leakage and coal wall splash were found.

In the process of advancing the working face for 100 m, the pressure fluctuation of the upper, middle and upper part of the bracket was more obvious, and the phenomenon of multiple cycles of pressure came. The pressure fluctuation in the lower part of the face was relatively stable, basically maintained between 16 and 22 MPa. The weakening of the roof made the average step of the cycle pressure to be controlled at 13 m.

7 Conclusion

  1. (1)

    Based on the specific geological conditions of 11,129 working face, the role relationship with the support bearing capacity of the roof during the initial and cyclic breaking process was firstly analyzed by using the thin plate theory of elastodynamics, and quantitative analysis was carried out by combined with different factors. It reveals that the high geological stress in the deep region and the characteristics of the rock layer itself (hardness and thickness) are the main controlling factors in determining the bearing ability of the support. The pressure within the controlled roof distance of the support is calculated to be in the range of 13.5–20.5 MPa.

  2. (2)

    The simulation part analyzes the decompression degree of roof stress before and after roof blasting, and the peak stress in front of the coal wall is reduced from 46.7 to 30.9 MPa, with obvious decompression effect. After roof blasting, the change of pressure value of different positions of the support slowed down, and the pressure value of the support at the working face ranged from 13.2 to 19 MPa, which was 33.6% lower than the pressure before roof blasting. The simulation results are similar to the theoretical results, which verifies the reliability of the simulation results.

  3. (3)

    The physical experiment part analyzes the mining instability and deformation behavior of the rock layer overlying the quarry after cutting the roof and its fracture evolution characteristics from the perspective of macroscopic mechanical manifestation, which is manifested as the overlying rock is destabilized by bottom-up mining break. The mining instability of the thick and hard bearing rock layer in the middle and low levels above the coal seam has the most obvious impact on the mining dynamic pressure on the support of the mining space and the coal rock body in front of it, while the mining instability of the hard bearing rock layer in the high level mainly affects the mining dynamic pressure on the crushed rock body accumulation in the hollow area behind the mining space. The peak pressure of the support during cyclic mining is in the range of 22.7–27.8 MPa, and the experimental results are closer to the theoretical calculations and simulation results.

  4. (4)

    The lithology of the roof of 11,129 working face and its characteristics were researched on site, and it was proposed to use the arrangement of "fan-shaped" blasting holes to carry out segmental blasting, with detailed parameters of the blast holes at different locations in the working face. The blasting effect is analyzed by the equipment on site, and the crack length can reflect the distance and intensity of the blast wave propagation inside the rock body. The rock mass at the beginning of the hole is highly broken, and the cyclic and strike cracks appear alternately in the middle and lower part of the drill hole, and the polygonal energy cuttings play a great effect. The support pressure during the cycle breaking after roof blasting is generally in the range of 18–30 MPa, and the artificial roof releasing is close to the theoretical analysis and simulation experiment results in a good improvement of the working environment of the quarry support, thus verifying the accuracy of the results of the analysis of a variety of means.