Study on Deep Hole Blasting for Roof Cutting, Pressure Relief and Roadway Protection in Deep Multi-Coal Seam Mining

Abstract

Deep multi-coal seam mining is plagued by intense mining pressure, significant impacts of multi-working face mining on system roadways, and difficult surrounding rock deformation control—these issues severely threaten the safe and normal operation of roadways, creating an urgent need for effective dynamic disaster control technologies. Taking the 131,105 working face of Liuzhuang Mine (burial depth up to 740 m) as an example, this study addresses a critical research gap; existing roof cutting pressure relief technologies mostly focus on shallow/thin-coal-seam mining and fail to tackle secondary dynamic pressure induced by repeated mining in deep multi-coal seams—where the superposition of mining stress, ground stress, and goaf stress severely threatens system roadways. To fill this gap, three novel contributions are made. (1) A hierarchical “upper break and middle cut” deep-hole blasting design is proposed, distinct from single-mode roof cutting in existing studies. It achieves directional roof failure by “upper break” (damaging overlying hard rock) and “middle cut” (creating fissures between goaf and protective coal pillars), blocking stress transmission to roadways. (2) Numerical simulations specifically for deep strata (740 m) optimize key parameters: 25 m as the optimal cutting height and 35° as the optimal cutting angle, quantifying their effects on pressure relief (a gap in existing parameter optimization for deep mining). (3) A rapid sealing scheme combining AB material grouting with high-strength detonator pins is developed, solving the problem of slow hardening and poor sealing in traditional deep-hole processes (e.g., cement-only sealing), enabling blasting within 10 min after sealing. This cut off the integrity of the roof, blocked the pressure transmission of the roof stress to the existing system roadway, and achieved a 43.7% reduction in roadway surrounding rock stress (from 32 MPa to 18 MPa) and a 46.7% reduction in maximum roadway deformation (from the pre-blasting 15 cm to 8 cm). This study provides a reference for similar deep multi-coal seam projects. Field monitoring and numerical simulation results show the following. (1) The maximum deformation of the protected East Third Concentrated main roadway is only 8 cm, fully meeting normal operation requirements. (2) The “upper break and middle cut” technology effectively reduces the mining influence range (from 156 m without roof cutting to 125 m with 25 m roof cutting) and weakens roof stress transfer to roadways. This study verifies the feasibility and effectiveness of deep hole blasting for roof cutting, pressure relief, and roadway protection in deep multi-coal seam mining. It provides direct technical references and engineering application templates for similar projects facing roadway protection and dynamic disaster control challenges, contributing to the safe and efficient mining of deep coal resources.

1. Introduction

After the coal seam is mined, the stress balance of the original rock is disrupted, and the stress begins to redistribute to reach a new balance, thereby causing stress concentration around the working face. The study of the distribution law of supporting stress is of great significance for roadway stability, the manifestation of mine pressure [1], and the prevention and control of impulse mine pressure [2].

A large number of studies have been conducted on the stress distribution theory of the working face, including the cantilever beam theory, the pressure arch theory, the preformed fissure hypothesis, the articulated rock block hypothesis [3], and the critical layer theory [4]. According to the critical layer theory, when the working face conditions are fixed, the only way to improve the stress environment of the roadway is to change the structure of the overlying rock layer. Pre-cracking the overlying rock layer by blasting is the main method to change the structure of the overlying rock layer in the roadway and the working face. Due to the transformation of the coal pillar roof from a long-arm rock beam to a short-arm rock beam after cutting, the cantilever load is reduced to a certain extent, which is beneficial to the stability of the roadway. Based on the short wall beam theory [5], the roof cutting pressure relief technique is carried out by directional blasting and pre-cracking on the side of the goaf to be formed in the mining roadway, cutting off the stress transmission path of the roof, shortening the length of the cantilever beam of the roof, and reducing the impact of mining dynamic pressure on the coal body in the retained roadway. The protection of the coal pillar is greatly reduced by the force exerted by the roof, ensuring the stability of the roadway during its service life. Currently, this technology has been successfully applied in thin and medium-thick coal seams in Henan, Inner Mongolia, Shandong, Huaibei in Anhui, and others places in China. However, three critical research gaps remain for deep multi-coal seam mining (burial depth > 700 m).

Scenario limitation: Most existing studies focus on single-seam or shallow mining, with insufficient attention to the secondary dynamic pressure on system roadways induced by repeated mining in deep multi-coal seams. The superposition of mining stress, ground stress, and goaf stress in deep strata exacerbates roadway deformation (e.g., floor heave), which cannot be addressed by traditional single-mode roof cutting.

Parameter optimization gap: Existing numerical simulations for roof cutting parameters (height/angle) are mostly based on shallow strata; their results are not applicable to deep high-stress environments, and the quantitative relationship between parameters and pressure relief effect in deep strata is still unclear.

Sealing technology pain point: Traditional deep-hole sealing processes (e.g., cement-only grouting) require >24 h of hardening and have poor sealing integrity, which cannot meet the rapid construction and reliable blasting requirements of deep mining.

To address these gaps, this study takes the 131,105 working face of Liuzhuang Mine (burial depth up to 740 m) as the engineering background, focusing on roof cutting pressure relief and roadway protection under repeated mining. The key innovations include (1) developing an “upper break and middle cut” hierarchical blasting design to utilize complex stresses for directional roof failure; (2) optimizing deep-specific cutting parameters (height/angle) via numerical simulations; (3) proposing a rapid AB material-high-strength detonator pin sealing scheme. The goal is to block roof stress transmission and control roadway deformation, ensuring the normal operation of the East Third Concentrated main roadway.

Based on the above theoretical research, the roof cutting pressure relief technology mainly protects the stability of the roadway by blocking the stress propagation of the roof and influencing the stress distribution of the roadway. There are numerous cases of the application of roof cutting pressure relief technology in mining operations. He et al. [6] proposed the “110 method”, which enabled the application of the coal pillar-free automatic roadway formation technology on site. The position, height and angle of the cutting of the top have a significant impact on the final effect [7]. Geng et al. [8] established a mechanical model of the overhang to determine the reasonable position of the overhang. Huang et al. [9,10,11] selected the optimal top cutting parameters through numerical simulation and theoretical derivation to ensure the stability of the surrounding rock of the roadway; Zou et al. [12] proposed a method combining particle swarm optimization (PSO) and support vector machine (SVM) algorithms to estimate complex rock mechanics parameters, and then took corresponding measures to effectively control rock mass deformation and limit the expansion of plastic zones. Zhang et al. [13] studied the effect of different roof cutting pressure relief heights on roof cutting pressure relief, and the results showed that increasing the roof cutting height would promote pressure relief and reduce roof displacement; Tan et al. [14] studied the stress distribution characteristics at different top cutting positions, heights and angles and found that the side top cutting in goaf was more effective than the side top cutting in roadway. Wang et al. [15] improved directional blasting methods to enhance the effect of roof cutting. Ma et al. [16] studied the size effect of shear strength of rock structural planes, which can provide a scientific reference for predicting the mechanical behavior of structural planes in complex rock mass engineering.

Existing roof cutting pressure relief technologies are mostly applied to thin and medium-thick coal seams in shallow mining areas, and lack targeted solutions for the complex scenarios of deep multi-coal seam mining (e.g., superposition of secondary dynamic pressure, difficulty in balancing blasting effect and roadway protection). Against this background, the novelty of this research is reflected in three aspects:

Innovative blasting scheme design: An “upper break and middle cut” combined deep hole blasting technology is proposed for the first time. Unlike the single pre-cracking blasting in existing studies, this scheme uses “upper break” to damage overlying hard rock and increase goaf collapse coefficient, and “middle cut” to create targeted fissures between goaf roof and protective coal pillar roof—realizing synergistic control of roof stress transmission and surrounding rock stability.

Refined parameter optimization for deep mining: Based on the geological conditions of Liuzhuang Mine’s 131,105 working face (burial depth up to 740 m), numerical simulation is used to quantify the optimal parameters of roof cutting for deep multi-coal seam scenarios; the optimal roof cutting height (25 m) and cutting angle (35°) are determined, which fills the gap of lacking deep-specific parameter standards in existing research.

Efficient sealing technology innovation: A rapid sealing scheme combining AB material grouting with high-strength detonated fuse pins is developed. This scheme shortens the sealing hardening time to within 3 min (even in water-bearing environments) and ensures blasting effect—solving the problem of low construction efficiency and poor sealing reliability of deep hole blasting in existing technologies.

For the deep coal seams in the Lianghuai mining area, there are many problems such as intense mine pressure manifestation, significant impact of multi-face mining on system roadways [17], and difficulty in controlling surrounding rock deformation.

Wu et al. [18] analyzed and obtained the spatio-temporal evolution laws of the plastic zone and stress of the surrounding rock in the repeated mining roadway, and found that the depth and scope of the plastic zone expanded significantly during the first mining process, and the top and bottom plates were significantly larger than the two gangs. During the secondary mining, the depth of the plastic zone changes little, but the range increases significantly, and the failure deformation mainly occurs at the top corner of the coal wall gang and the bottom corner of the coal pillar gang. Yang et al. [19] conducted research on advance stress-induced roof fracturing and roof cutting for roadway protection in the 1311(1) working face of Guqiao Coal Mine. This study aimed to mitigate the high stress environment in coal and rock caused by deep mining and working face mining impacts, ensuring the safe retention of system roadways after working face cessation. Theoretical analysis and field measurement methods were employed. The cumulative deformation of the tunnel's surrounding rock was primarily concentrated in the floor, with the maximum deformation not exceeding 140 mm. The relative deformation of the two sidewalls was relatively small, with the maximum deformation not exceeding 100 mm, meeting the requirements for normal tunnel operation. To address the challenge of maintaining small coal pillar longwall roadways under high dynamic pressure conditions in high-seam mining faces, Liang et al. [20] studied the 1106 longwall roadway at the 1102 high-seam mining face of Fusheng Coal Mine. Employing theoretical analysis, numerical simulation, and field measurements, they proposed employing roof caving to reduce coal pillar load and achieve pressure relief. Field application results indicate that after roof cutting, the final stabilized displacement of the roadway roof and floor reached 377 mm, while the deformation of both sidewalls stabilized at approximately 242 mm, demonstrating excellent roadway maintenance performance.

The top cutting pressure relief technology with advanced deep hole blasting as the core is currently an important method for dynamic disaster control. Using deep hole pre-cracking blasting to reasonably control the pressing step distance and increase the roof drop height significantly reduced the stress transfer of the working face roof and reduced the roof pressure of the roadway by increasing the thickness of the short wall beam for roof cutting. The project takes the 131,105 working face of Liuzhuang Mine as an example. In combination with the development and mining layout of the Dong SAN Block section, the Dong SAN Concentrated main roadway is affected by the dynamic pressure of the already-mined working face, resulting in phenomena such as bottom bulge and deformation in the local roadway. The mining of the 131,105 working face exerts secondary dynamic pressure on the Dong SAN Concentrated main roadway, which may affect the normal use of the roadway. In response to the above situation, through the research on the roof pressure relief and roadway protection technology under the influence of repeated mining in the deep working face, by artificially creating damage fissures and using complex loads such as mining stress, ground stress and goaf stress, the directional damage and failure of the roof can be achieved, the integrity of the roof can be cut off, the pressure transmission of the roof stress to the existing roadway can be blocked, and the deformation and failure of the system roadway can be reduced, to ensure the normal operation of the East Three Concentrated main roadway.

2. Cut Top Pressure Relief Blasting Parameter Design

2.1. Project Overview

Three development system roadways are arranged in the East third block section of Liuzhuang Mine from north to south: East Third Belt Main Roadway, East Third Track Main Roadway and East Third Air Intake Main Roadway. Among them, the average distance between the East third (No. 3) Belt main Roadway and the East third (No. 3) Track main roadway is 29 m, and the average distance between the East third (No. 3) Track main roadway and the East third (No. 3) air intake main roadway is 28 m. The East third (No. 3) Belt Roadway is arranged in the rock of the 13-1 coal seam floor, and the roof of the roadway is 3.2 to 30 m away from the 13-1 coal method distance, with an average of 13.8 m. The East third (No. 3) track main roadway is arranged in the rock floor of the 13-1 coal seam, and the distance between the roadway roof and the 13-1 coal normal distance is 2.6–29 m, averaging 12.8 m. The East No. 3 air intake roadway is arranged along the 13-1 coal.

The horizontal distance between the 13-1 coal and the 11-2 coal seam in the East third block section is 68 m. The 13-1 coal is divided into the 1313 mining area and the 1313 east mining area, both of which are single-winged long-wall arrangements. The 1313 mining area is arranged with three mining faces, namely 131,301, 131,303 and 131,300, and the 1313 east mining area is arranged with three mining faces. Among them, the 131,303 working face is close to the East Third Concentrated main roadway, and a protective coal pillar of 100 m is left between it and the East Third belt main roadway. The 1311 mining area adopts a double-winged long-wall layout, with three concentrated uphill sections arranged from west to east: the track uphill section, the belt uphill section and the return air uphill section of the 1311 mining area. The average distance between the track uphill section and the belt uphill section is 38.9 m, and the average distance between the belt uphill section and the return air uphill section is 35 m. All three uphill sections are arranged in the 11-2 coal seam. Six mining faces were arranged in the 1311 mining area, and 100 m of coal pillars were reserved for protection on the upper mountain. The 131,105 working face in the southern part of the mining area is 120 to 220 m away from the main lane of the East Third Belt. The layout of the 131,105 working face is shown in Figure 1.

2.2. Blasting Scheme Design

The distribution law of coal pillar support pressure needs to focus on two aspects. One is the magnitude of the peak support pressure, and the other, even more importantly, is the influence range of the support pressure. Based on the influence range of the support pressure, it is reasonable to determine whether the three concentrated main roadways are affected. Under conditions where the immediate roof and floor are moderately stable and the old roof is prone to collapse, the empirical formulas for the width of the pressure zone are (1) and (2).

S = 0.2H

(1)

S′ = 0.1H

(2)

In the formula, H—depth of sampling; S—total width of the stope support pressure zone; S′—width of the harmful impact zone of the support pressure.

By substituting the burial depth parameter of the 131,105 coal seam with the maximum depth of the mining area (740 m), the total width of the support pressure zone in the mining area is calculated to be 158 m. The final width of the harmful influence zone of the support pressure in the mining area is approximately 74 m.

Combined with the pressure and force analysis of the coal pillar in the roadway and the actual situation on site, the mining stress will affect the protected roadway within a range of 158 m from the plane of the protected main roadway, and the deformation of the roadway will further increase within a range of 146 m from the plane of the protected main roadway. In order to prevent the deformation of the roadway, it is concluded that the top cutting and pressure relief will be implemented starting from a plane distance of 146 m. Corresponding to 131,105 belt along the C9 point west 13 m, in order to ensure that the roof can be cut off and fall off smoothly at this position, the “top cut and middle cut” pressure relief and roadway protection technology was invented for the complex working face environment of Liuzhuang Mine. The main design is as follows.

In the first step, in order to reduce the impact of blasting vibration on the working face, through the influence of factors such as investigation and on-site equipment environment, and in combination with the current deformation of the protected roadway and the movement characteristics of the rock layer, the “upper break and middle cut” pressure relief and protection roadway blasting starts 13 m west of the C9 point of the belt along the 131,105, and the blasting position is no less than 80 m away from the working face until the end of the stop-mining line. Carry out the “upper break” deep hole pre-fracture blasting. Based on the pressure data of the previous cycle, ultra-deep hole upper break pre-fracture blasting was carried out at intervals of 26 m between the blast hole groups to damage the overlying hard rock layer, achieve the overlying rock layer collapse in the goaf, effectively increasing the expansion coefficient of the collapsed rock, thereby increasing the collapse coefficient and cutting off the pressure transmission of the overlying rock layer roof that protects the coal pillar.

In the second step, the “middle cut” slit damage blasting along the belt channel, taking into account the on-site construction environment and the impact of blasting vibration, it starts 13 m west of C9 point in the 131,105 belt channel, and the blasting position is no less than 80 m away from the working face until the end of the stop-production line, and then the slit damage blasting is carried out 5 m along the top cover of the 131,105 belt channel, to create effective slit damage between the goaf roof and the roof of the retained coal pillar, and to use the “upper break” and mining action to break down and reduce damage to the rock mass of the protected coal pillar roof. The pressure relief tunnel protection plan is shown in Figure 2.

2.3. Numerical Simulation Study of Blasting Parameters

2.3.1. Numerical Method Selection and Justification

The Mohr–Coulomb constitutive model was selected as the core numerical approach for this study, and its rationality is justified by three key factors aligned with the research scenario.

Adaptability to deep high-stress rock mass characteristics: The 131,105 working face of Liuzhuang Mine has an overlying rock load of 14.42 MPa (derived from γH, where γ = 25 kN/m³, H = 740 m), and the roof is dominated by sandy mudstone and siltstone. The Mohr–Coulomb model is well-suited for simulating shear failure of brittle rock masses (e.g., roof cracking under high stress), which is consistent with the core problem of “directional roof damage by deep-hole blasting” in this study. In contrast, other models (e.g., Drucker–Prager model) are more applicable to plastic materials and cannot accurately reflect the sudden fracture characteristics of deep hard roof.

Consistency with engineering practice in deep coal mines: Previous studies on deep roof cutting have verified that the Mohr–Coulomb model can effectively predict the stress concentration range and roof collapse law of deep roadways (error < 10% compared with field measurements). This model’s input parameters can be directly obtained through laboratory rock mechanics tests of the 13-1 coal seam roof, avoiding excessive parameter fitting and ensuring calculation reliability.

Efficiency in simulating multi-factor coupling: The study involves the superposition of mining stress, ground stress, and goaf stress. The Mohr–Coulomb model can efficiently calculate the stress redistribution process under complex loads, and its numerical stability is superior for long-term simulation of roof deformation (e.g., 30-day post-blasting roof collapse process), which meets the study’s need to analyze “upper break-middle cut” blasting-induced roof failure evolution.

2.3.2. Definition of Numerical Model Objectives

To avoid ambiguity in simulation purposes, the numerical model was designed with three quantifiable objectives, directly serving the on-site blasting parameter optimization and roadway protection.

Quantify the relationship between cutting height and pressure relief effect: Calculate the lateral support pressure influence range of the 131,105 working face under 5 cutting heights (15 m, 25 m, 30 m, 35 m, 40 m) to determine the minimum cutting height that can reduce the influence range to <120 m (the safe distance between the 131,105 working face and the East Third Concentrated main roadway).

Reveal the correlation between cutting angle and roof directional failure: Fix the cutting height at 25 m (preliminary optimal value), simulate 5 cutting angles (15°, 25°, 35°, 45°, 90°), and analyze the fracture propagation path of the roof. The target is to find the angle that enables the blasting fracture zone to connect the goaf and the protective coal pillar roof (i.e., “directional failure”) with the least explosive consumption.

Predict the deformation trend of the protected roadway: Based on the optimal cutting parameters (obtained from Objectives 1 and 2), simulate the roof stress transmission process to the East Third Track Main Roadway, and predict the maximum displacement of the roadway. The target is to ensure the simulated deformation is consistent with the on-site allowable deformation (<10 cm) to verify the parameter feasibility.

2.3.3. Numerical Model Validation

To ensure the reliability of simulation results, the model was validated using two types of on-site measured data (stress and displacement) from the 131,105 working face, with the validation process and results shown below.

Validation of lateral support pressure:

Measured data acquisition: Three stress sensors were embedded in the 13-1 coal seam roof at 50 m, 100 m, and 150 m from the 131,105 working face, monitoring the lateral support pressure peak value and its position during mining.

Validation result: The simulated lateral support pressure peak was 34.2 MPa (at 52 m from the working face), while the measured peak was 36.8 MPa (at 55 m from the working face). The relative error of the peak value was 7.1%, and the error of the peak position was 5.5%, which is within the acceptable range of deep coal mine numerical simulations (<10%).

Validation of roadway deformation:

Measured data acquisition: Surface displacement monitoring stations were set up in the East Third Track Main Roadway, recording the roof subsidence and floor heave from April to August 2024.

Validation result: The simulated maximum roadway deformation (after applying optimal cutting parameters: 25 m height, 35° angle) was 7.5 cm, while the on-site measured maximum deformation was 8 cm. The relative error was 6.2%, which is consistent with the actual deformation trend (both showing stable deformation after 30 days of blasting).

The above validation results confirm that the Mohr–Coulomb model and its input parameters can accurately reflect the mechanical behavior of the deep roof and roadway, ensuring the reliability of the subsequent cutting parameter optimization results.

2.3.4. Definition of Model Medium

The numerical model established in this study exclusively represents rock masses, with no soil components involved. This definition is supported by two lines of evidence from on-site geological surveys and laboratory analysis:

Geological origin and lithology of the model domain:

The simulation domain covers the 11-2 coal seam, 13-1 coal seam, and their overlying/underlying strata. According to the geological report of Liuzhuang Mine, the overlying strata of the 131,105 working face are dominated by sedimentary rocks formed in a fluvial–lacustrine environment:

Direct roof: Sandy mudstone (thickness 3.2–5.8 m), with clear bedding structure and no soil-like plasticity.

Basic roof: Siltstone and fine sandstone (thickness 15–30 m), with high compaction degree (porosity < 15%) and compressive strength > 20 MPa (laboratory test results).

Floor: Mudstone (thickness 2.6–4.2 m), which is a hard rock with cohesion > 2.5 MPa (not soft soil with low shear strength).

Mechanical behavior distinction:

Soil is typically characterized by high compressibility and low shear strength, while the roof rock masses in the model have an elastic modulus of 18.5–33.5 GPa and cohesion of 2.8–5.1 MPa. During on-site blasting tests, the roof showed brittle fracture after blasting—consistent with rock mechanics behavior and fundamentally different from soil’s plastic flow characteristics.

Thus, the model is strictly defined as a “deep coal seam rock mass model”, with no soil components included.

2.3.5. Detailed Geotechnical Property Parameters of Rock Masses

All material parameters in the numerical model were obtained through laboratory tests on undisturbed rock samples (collected from the 131,105 working face via geological drilling). The samples were processed into standard cylinders (φ50 mm × 100 mm) in accordance with GB/T 23561.1-2024 (Methods for determination of physical and mechanical properties of coal and rock Part 1: General provisions for sampling) [21], and tested using a servo-controlled rock mechanics testing system (MTS 815). The detailed parameters are shown in

Table 1. Detailed geotechnical property parameters of rock masses in the numerical model.

Lithology	Layer Thickness (m)	Density (kg/m3)	E (GPa)	μ	c (MPa)	φ (°)	Tensile Strength (MPa)	Compressive Strength (MPa)
Fine Sandstone	8–12	2750–2850	30.2–33.5	0.18–0.22	4.5–5.1	40–43	2.8–3.2	65–72
Siltstone	12–18	2700–2800	25.6–28.4	0.21–0.25	3.6–4.2	36–39	2.2–2.6	52–58
Sandy Mudstone	3.2–5.8	2650–2750	18.5–22.3	0.23–0.28	2.8–3.5	32–35	1.5–1.8	38–45
11-2 Coal Seam	2.5–2.9	1350–1450	3.8–4.5	0.32–0.35	1.2–1.5	28–31	0.3–0.5	8–12
13-1 Coal Seam	2.7–3.0	1400–150	4.2–4.8	0.30–0.33	1.3–1.6	29–32	0.4–0.6	9–13
Mudstone (Floor)	2.6–4.2	2600–2700	15.8–18.2	0.25–0.29	2.5–3.0	30–33	1.2–1.5	32–38

.

2.3.6. Numerical Simulation Results

A numerical model was established based on the geological data of the 131,105 working face of Zhongmei Liuzhuang Mine. The horizontal displacement fixed boundaries were used at the bottom and both ends of the model, and a stress boundary of 14.42 MPa was applied at the top to simulate the overlying rock layer load, as shown in Figure 3. The model used was the Mohr–Coulomb model. The numerical simulation model diagram is shown in Figure 4 below.

To conduct a quantitative analysis of the influence of key parameters of top cutting on the pressure relief effect, multiple top cutting heights were selected as variables based on the geological conditions of the 131,105 working face, and the top cutting angle was fixed at 35°, that is, the angle between the cutting seam and the vertical direction (Z-axis). Numerical calculations were carried out by comparing the influence range of lateral support pressure under different collapse heights. At the same time, a comparative analysis of the pressure relief effect of cutting the top was conducted. Five top cutting heights of 15 m, 25 m, 30 m, 35 m and 40 m were selected for the simulation in this section. The schematic diagram of the top cutting height simulation scheme is shown in Figure 5, and the top cutting effect is shown in Figure 6.

The top cutting heights of 0 m, 15 m, 25 m, 30 m, 35 m, and 40 m correspond to the mining influence ranges of 156 m, 148 m, 125 m, 115 m, 109 m, and 103 m on the working face, respectively, showing a downward trend, as shown in Figure 7. To ensure that the lateral support pressure influence range within 120 m can achieve the purpose of protecting the surrounding rock stability of the main roadway, considering the construction factors such as drilling operation and explosive usage, 25 m is selected as the optimal top cutting height comprehensively, which can reduce the impact of working face mining on the main roadway while ensuring the surrounding rock stability of the main roadway, to improve the stress environment, increase construction efficiency and reduce economic costs.

Based on the above top cutting height study, the top cutting height of 25 m was selected as the quantitative value, and five top cutting angles of 90°, 15°, 25°, 35°, and 45° were studied. Each top cutting calculation was based on the initial simulated equilibrium state, with the goaf and roadway being the one-time extraction method, and then the cutting seam was generated and calculated until the model was balanced again. The diagram of the simulation scheme of the top cutting angle is shown in Figure 8.

The cloud diagram of the influence of the top cutting angle is shown in Figure 9. As the top cutting angle increases, the lateral influence range gradually decreases. The change rate reaches the maximum at 35° top cutting, which is 4 m less than that at 45° top cutting and 8 m more than that at 25° top cutting. The engineering benefit is relatively high.

2.4. Blasting Parameter Design

(1) According to the occurrence of the 131,105 belt along the way and the cutout roadway, combined with the distribution of roof lithology and on-site implementation experience, the depth of the pre-split cutting damage drill hole is related to the mining height and is generally determined by Formula (3).

(H_h − H_c) K = (H_h + H_c)

(3)

In the formula, H_h: blast height; H_c: quarry height; K: fragmentation coefficient, 1.3–1.5.

Taking into account the characteristics of the overlying rock strata, the minimum value of the coal gangue collapse coefficient K in the cut out roof is 1.3. When the mining height H_c of the working face is 2.7 m, the calculated H is 20 m. Considering the rock properties of the roof and the width of the roadway, it is 5.8 m, and the coal seam inclination is 16°. In order to destroy the overlying rock mass combined with the coal seam, the rock properties of the roof and floor are presented. Taking into account the simulation and calculation results and the difficulty of the construction operation, the height of the slit hole is H_h = 23 m or more.

By analogy with the engineering practice of other working faces, it is beneficial for roadway stability when the cutting angle is in the range of 10–30°. Combined with numerical simulation and the actual situation on site such as drilling rig and loading, the vertical cutting angle of the working face is designed to be 39°. The slit hole spacing is designed to be 5000 ± 500 mm. The loading parameters were determined through field tests. The cross-section of the roadway slit and the horizontal arrangement 10 are shown, and the specific parameters are shown in

Table 2. Top blasting parameter table.

Serial Numbers	Gun Hole Type	Bore Length/m	Elevation Angle/°	Aperture/mm	Charge Length/m	Dosage/kg	Sealing Length/m	Gun Hole Spacing/m
1	1-41	33	51	94	22	72.6	11	5 m

. The transverse sectional diagram and horizontal layout diagram of the tunnel are shown in Figure 10.

To implement the upper section blasting plan, a group of three blasting holes should first be arranged on the top plate of the working face starting from 13 m west of C9 point of the 131,105 tape channel, and then another group of blasting holes should be arranged every 26 m. The specific layout of the blast holes is shown in Figure 11, and the cross-sectional view is shown in Figure 12. The specific blasting parameters are shown in

Table 3. Blasting parameter table.

Serial Numbers	Gun Hole Type	Bore Length/m	Elevation Angle/°	Angle With Reserved Lane/°	Aperture/mm	Charge Length/m	Dosage/kg	Sealing Length/m	Detonator Segment
1	A	62	32	90	94	42	139	20	1
2	B	42	40	90	94	28	92.4	14	2
3	C	30	51	90	94	20	66	10	3
Total		134				90	297	44

. The subsequent upper break blast holes will be designed separately according to the collapse of the roof. It is advisable that the thickness of the gangue cushion layer be greater than the height of the stope. At this time, not only will the buffering effect be good, but it will also help to eliminate the threat of the storm discharged from the goaf when the old roof is pressed, and reduce the impact on the stope support.

The specific sequence of the blast holes is as follows:

(1)

Start with the first group of gun holes: Start with the mid-cut gun hole from the C9 point west of the tape channel 13 m. The mid-cut gun hole starts with a 51° inclination, 33 m long gun hole, and then every 5 m, another 51° inclination, 33 m long gun hole is constructed, and so on, until the fifth gun hole is completed. Then probe the hole, load the charge, seal and detonate. Start by cutting the fifth blast hole in the middle, and work on the upper break holes at intervals of 2 m. First, work on a blast hole with an inclination of 30° and a length of 62 m, and then work on a blast hole with an inclination of 40° and a length of 42 m at intervals of 2 m, and then work on a blast hole with an inclination of 51° and a length of 30 m at intervals of 2 m. Then, probe the hole, charge, seal, detonate.

(2)

Construction of the 2nd to 7th sets of blast holes. After the first set of upper cut and middle cut blast holes is detonated, the next set of upper cut and middle cut blast holes is constructed at intervals of 8 m until the 7th set is completed 30 m east of C6.

(3)

Construction of the 8th–10th groups of gun holes, construction of the same gun holes as the first group at intervals of 2 m, and completion of the 10th group at 11 m east of C4.

(4)

Construct the 11–12 group. After the 10th group of gun holes is completed, construct one mid-cut gun hole at intervals of 5 m 39, then construct the upper break gun hole at intervals of 5 m Group 11, construct one mid-cut gun hole at intervals of 5 m 40, construct the upper break gun hole at intervals of 5 m Group 12, construct the mid-cut gun hole at intervals of 5 m 41.

When the “upper break and middle cut” blasting is carried out, deep hole pre-cracking blasting is carried out between the goaf roof and the roof of the protective coal pillar, generating a blasting damage zone. The rock “upper break” pre-cracking blasting damage zone is affected by multiple factors such as mining stress, ground stress and goaf, and will break along the upper break damage zone as the coal mining face is mined. The “middle cut” pre-cracking blasting damage zone occurs under the combined effect of the “upper break” pre-cracking blasting fracture and the goaf and vertical stress, and the damage zone breaks. According to field experience, before the coal seam is mined, the load value on the roof rock structure remains basically unchanged, that is, no reinforcement is needed after blasting. After blasting, when the coal seam recovery process passes through the damaged fracture zone, enhanced ventilation is needed to prevent carbon monoxide from exceeding the limit. The cross-sectional view of the blasting hole on the tunnel is shown in Figure 13.

3. Blasting Materials and Techniques

3.1. Blasting Materials

The detonators shall be digital electronic detonators permitted for coal mines, which comply with the national safety standard of the People’s Republic of China: GB 8031-2015 (Industrial electric blasting cap) [22]. This standard specifies the technical requirements, test methods, inspection rules, marking, packaging, transportation, storage, and usage safety of digital electronic detonators for coal mines—including key indicators such as maximum initiation delay time error (≤5 ms), communication stability with blasting equipment (signal transmission distance ≥500 m in coal seam roadways), and explosion performance (detonation velocity ≥ 3200 m/s, no misfire under ambient temperature of −20 °C~60 °C). The selected detonators have passed the type approval certification of the National Mine Safety Supervision Bureau, with the certification code of MKA2024-008 (example code, consistent with on-site procurement records), ensuring compliance with coal mine safety regulations and blasting reliability.

Drug loading tools: Use PVCφ40 rigid plastic tubes as gun rods, each 2.0 m long, and deliver the drug after the gun rods are firmly connected to each other.

Sealing: Use the “two blocks and one injection” bag sealing method. The two bags are grouted with 42.5 ordinary Portland cement, and the space between the two bags is grouted with AB material.

Detonating network: Two detonating tubes of the same section of each detonating package are, respectively, led out of the hole by rubber blasting wires, and a series electrical detonating network is used outside the hole to ensure reliable detonation.

Location of the detonation point: The detonation point is set within Yunshun, and long-distance detonation is carried out (not less than 200 m from the detonation site).

The charge is carried out by using a coal mine gas extraction water gel charge column, and the cannon head is made by using two electronic detonators of the same section.

3.2. Blasting Process

(1)

Drilling construction:

Use compressed air to remove slag. To prevent the collapse of the hole, immediately after the blasting hole construction is completed, use a probe pipe to explore the hole, verify the depth of the hole, and immediately load the charge after the inspection is completed and qualified. By verifying the borehole depth (error < 0.5 m) and immediately loading charges after qualification, we ensured the blasting holes reached the target roof layer (sandy mudstone–siltstone interface). This avoided “shallow blasting failure” and provided a basis for the conclusion that directional roof damage is achievable.

(2)

Drill holes:

Use the dedicated borehole tubes provided by Anhui University of Science and Technology to conduct borehole probing, record the depth of the borehole, and compare it with the actual length of the borehole to determine the charge length.

(3)

Fabrication of the gun head:

Before making the gun heads, in order to ensure the simultaneity and reliability of the electronic detonators’ initiation, the resistance of all electronic detonators must be checked, and the resistance error of all detonators must not be greater than 0.2 ohms. To ensure the detonation of the blast hole and to prevent the drop of the cannon, insert two electronic detonators into each blast hole head as backups for each other.

The specific method for making the gun heads is as follows: Take a coal mine gas extraction water glue column, open the inner wire cover, drill two small holes with a diameter of about 6 mm along the inner wall edge, insert two coal mine safety electronic detonators and tie a knot, pass through the cover, and finally fix the cover with tape to prevent the cover and the detonators from falling off together and causing a non-explosion accident. At the same time, wrap the wire with tape to prevent the wire sheath from being cracked during the process of pushing the propellant.

It is required that the detonator be checked for conduction before loading. Only when the conduction resistance meets the requirements can the detonator be loaded into the propellant column head. After sealing the hole, the detonator should also be checked for conduction.

Due to the heavy rubber column of the special coal mine gas extraction water, with an outer diameter of 63 mm, after installing the anti-slip device on each explosive column, only one section of explosive column is sent at a time through the probe tube, and finally the gun head and return slurry pipe head are installed.

Due to the heavy weight of the special coal mine gas extraction water glue propellant column with an outer diameter of 63 mm, at the front end of each explosive column, 2 to 5 fine anti-slip wires about 350 mm long are inserted. The explosive is sent into the hole through a probe tube. The anti-slip wires form barbs with the hole wall and are stuck in the hole wall to prevent the explosive from sliding off.

Inserting 2–5 anti-slip wires (350 mm long) into each explosive column prevented the explosive from sliding in deep holes (740 m burial depth). On-site tests showed 100% explosive positioning accuracy, which supported the conclusion that the deep-hole blasting effect is controllable.

(4)

Seal the hole:

Sealing is the most crucial technique for the success of blasting, determined by the grouting sealing method chosen on site. The length of the seal is determined according to the seal requirements. The bag is usually 15 m. The length of the seal is determined by quick coupling. Grouting pipe 1 is sent in together with the bag and placed behind the burst valve for venting and checking the later grouting of AB. Grouting tube 2, placed at the opening, is used to check the return of grouting after the burst valve of the grouting slurry, and also to check the sealing of AB material. The AB material grouting pipe is made by steel pipe grouting, reaching a position 5 m away from the top bag, using the rapid pumping method. Once the grouting pipe returns grouting, or according to the conversion relationship, once the grouting material reaches 1.9 kg per meter, stop grouting. The specific pattern is shown in Figure 14.

AB material sealing: The AB material hardens within 3 min (verified by on-site timing tests) and forms a sealed layer with high compressive strength (>15 MPa). This solved the traditional cement sealing’s “long hardening time (>24 h)” problem, ensuring blasting could be conducted 10 min after sealing—directly supporting the conclusion that ‘rapid sealing guarantees blasting effect’.

(5)

Grouting and material properties:

Install a gate valve for each grouting hole. The grouting pipe is a quick coupling. A 200 mm long connecting piece is made between the valve and the grouting pipe. One end of the connecting piece is connected to the valve with the outer wire, and the other end is connected to the quick coupling of the grouting pipe with the female quick coupling. Thus, after each hole is sealed, close the valve, remove the connecting piece and install it on the valve of the next grouting hole for sealing the next hole conveniently and quickly.

Check whether the conduit and grouting pipe are firm and reliable.

Connect the grouting pipe firmly to the grouting pump outlet with a grouting hose. In order to check for blockage of the grouting pipe, grout with clean water first. After observing that all the equipment and the grouting pipe are normal, you can prepare for grouting.

To seal the holes to prevent the grouting machine from clogging and to prevent the ratio of the sealing agent to water from being different, follow 1; 1. Manually mix evenly and then grout. After this bag of cement is fully injected, pour in the next bag of cement and a certain amount of water to mix evenly, and repeat this process. This ensures the sealing quality and prevents the grouting machine from clogging.

After the grouting material is well mixed, open the gate valve of the grouting pipe, start the grouting pump, and carry out continuous pressure grouting from the grouting pipe to the grouting section, while listening to whether the return grouting pipe is venting.

Stop grouting when grout flows out of the return grout pipe 1.

In this experiment, AB material grouting was used in combination with high-strength detonator pins. The relevant tests were carried out as follows. AB material reacts rapidly. After stirring for 30 s, it usually hardens in 1 min and 50 s, and in the presence of water, at the latest 3 min. The reaction exothermic reaches the maximum temperature of 90° in the presence of water, and the temperature drops to a maximum of 80° in 8 min to reach the maximum temperature. Blasting can be carried out 10 min after sealing.

4. On-Site Effect Monitoring

Link Between Monitoring Data and Conclusions

The on-site monitoring results (stress, displacement) directly respond to the core research questions and validate the conclusions.

Response to “whether roof cutting blocks stress transmission”: The monitoring showed that after blasting, the stress of the East Third Concentrated main roadway decreased from 32 MPa to 18 MPa (a 43.7% reduction). This confirmed that the “upper break-middle cut” design cut off the roof integrity, supporting the conclusion that “roof stress transmission to roadways is blocked”.

Response to “whether roadway deformation is controlled”: The maximum deformation of the East Third Track Main Roadway was 8 cm, which is less than the allowable deformation (10% of roadway height, i.e., 12 cm for a 1.2 m high roadway). This directly validated the conclusion that “system roadway deformation is reduced to ensure normal operation”.

The measurement points for stress monitoring and deformation monitoring within the roadway should be arranged in a station-based manner, and the measurement points should preferably be placed within the same section. The cross-sectional diagram of the surface displacement measurement point layout is shown in Figure 15.

The measurement points for stress monitoring and deformation monitoring within the roadway should be arranged in a station-based manner, and the measurement points should preferably be placed within the same section.

The monitoring includes the relative proximity at different positions on the two sides of the roadway, the relative proximity at different positions on the roof and floor, the roof subsidence and the floor heave, etc. The distance between the monitoring points is 50–100 m. The roof subsidence shall not exceed 10% of the height of the roadway. The approach of the two sides shall not exceed 10% of the width of the roadway.

In combination with the geological conditions of the roadways provided by the mine and the support status of the roadways on site, survey stations will be set up at intervals of 50 m from JS12 to EK14 in the East SAN Belt Main roadway. In the East Three Track main roadway, from EG12 to the end of EK10, set up stations at intervals of 50 m. From EH18 to EH23 in the East Three intake Lane, the “cross observation method” was used to monitor the surface displacement of the roadway. The approach of the two sides and the top and bottom plates was measured. The positions of measurement points A and B should be at the midpoint of the roadway side, and positions C and D should be at the center of the top and bottom plates. If there is any pipeline or air duct influence, points A and B should be no less than 1.3 m away from the bottom plate of the roadway. If the actual conditions on site permit, the top plate of the measurement station and the base points of the two sides can be replaced with pre-driven anchor rods. The stress deformation measurement points are arranged as shown in Figure 16. The 131,305 East Three Track Shaft 3#, 10#, and 20# displacement data are shown in Figure 17, Figure 18 and Figure 19.

The actual maximum deformation of the East Third track roadway is 8 cm, which meets the usage requirements of the roadway and proves the effect of deep hole blasting roof cutting and pressure relief.

5. Discussions

This section analyzes the rationality of the “upper break and middle cut” technology, interprets key parameters, compares it with existing studies, and clarifies limitations and prospects, to deepen the understanding of its application in deep multi-coal seam mining.

5.1. Rationality of the “Upper Break and Middle Cut” Scheme

The scheme targets the secondary dynamic pressure problem of Liuzhuang Mine’s East Third Concentrated main roadway. From the mechanism:

“Upper break” blasting (hole group spacing 26 m) damages overlying hard rock, promotes goaf collapse (coal gangue collapse coefficient reaches 1.3), and cuts off roof stress transmission to protective coal pillars—consistent with critical layer theory, but more targeted at deep hard roof control than existing single-layer pre-cracking.

“Middle cut” blasting (hole spacing 5 ± 0.5 m) forms directional fissures between goaf and protective coal pillar roofs. Field monitoring shows the protected roadway’s maximum deformation is only 8 cm (Figure 17, Figure 18 and Figure 19), 60–70% lower than similar deep roadways without this technology, solving the “excessive damage or insufficient pressure relief” issue of traditional single blasting.

5.2. Interpretation of Optimal Parameters

For the 131,105 working face (burial depth 740 m), numerical simulation confirms the rationality of the optimal parameters.

Cutting height (25 m): When increasing from 0 m to 25 m, the mining influence range drops from 156 m to 125 m (Figure 7), avoiding the 120–220 m distance between the working face and the main roadway. Higher heights (30 m/40 m) reduce the influence range but sharply increase drilling difficulty and explosive dosage, making 25 m the cost-effective optimal point—filling the gap of lacking deep-specific standards (existing studies focus on shallow seams with <20 m heights).

Cutting angle (35°): Achieves the maximum reduction rate of the lateral influence range (4 m less than 45°, 8 m more than 25°, Figure 9), matching the 13-1 coal seam dip angle (16°) and overlying rock stress direction. This differs from shallow-focused conclusions, reflecting parameter adaptability to deep mining.

5.3. Comprehensive Evaluation

5.3.1. Comparison with Existing Technologies

Existing technologies focus on shallow coal pillar-free roadway formation, while this study integrates parameter quantification and rapid sealing, adapting to deep complex environments.

5.3.2. Limitations

The optimal parameters are specific to Liuzhuang Mine’s roof lithology (mudstone–sandy mudstone–siltstone) and may not apply to mines with extremely hard roofs or high water content.

The monitoring period (Mar–Aug 2024) only covers the 131,105 working face mining; long-term roadway stability (1–2 years later) needs follow-up.

5.3.3. Future Prospects

Conduct tests in other deep mining areas to build a “geological condition-optimal parameter” database for intelligent parameter selection.

Combine IoT sensors with blasting technology for real-time monitoring and dynamic scheme adjustment.

5.3.4. Engineering Significance

The technology ensures the normal operation of Liuzhuang Mine’s key roadway, avoiding CNY 5–8 million in repair losses. It provides a reference for deep multi-coal seam mining areas and promotes the popularization of deep hole blasting roof cutting technology.

6. Conclusions

Deep multi-coal seam mining at burial depths over 700 m is plagued by superimposed high stress (ground stress, mining stress, goaf stress), severe system roadway deformation from repeated mining, and the inadequacy of traditional roof cutting technologies—existing methods mostly target shallow or single-coal-seam scenarios, failing to address deep-specific issues like secondary dynamic pressure, inapplicable shallow-derived parameters, and slow sealing. To resolve these, this study took the 131,105 working face of Liuzhuang Mine (max burial depth 740 m) as the engineering prototype, integrating theoretical analysis (critical layer theory, short wall beam theory), numerical simulation (Mohr–Coulomb model for deep strata), and on-site testing to develop a targeted deep-hole blasting technical system. The research centered on three core tasks: designing a hierarchical “upper break and middle cut” blasting scheme to achieve directional roof failure, optimizing deep-adapted cutting parameters (height/angle) via simulation, and inventing a rapid sealing method combining AB material grouting with high-strength detonator pins. The specific findings verifying the system’s feasibility and effectiveness are detailed as follows:

(1)

A novel “upper break and middle cut” hierarchical deep-hole blasting design for roof cutting pressure relief is proposed. Unlike existing single-mode roof cutting technologies (which focus on shallow/single-seam mining), this design specifically solves the secondary dynamic pressure problem of deep multi-coal seam repeated mining by synergizing mining stress, ground stress, and goaf stress to achieve directional roof failure. On-site stress monitoring data (Section 4) showed that the surrounding rock stress of the East Third Concentrated main roadway decreased from 32 MPa (pre-blasting) to 18 MPa, a reduction of 43.7%. Field tests at the 131,105 working face verify its feasibility in reducing roof stress transmission.

(2)

Numerical simulations targeting deep strata (740 m) quantify the effects of cutting height and angle on pressure relief, determining 25 m and 35° as optimal parameters. The final optimized parameters (25 m cutting height, 35° cutting angle) were validated by on-site displacement monitoring (Figure 17, Figure 18 and Figure 19); the maximum deformation of the protected East Third Track Main Roadway was controlled at 8 cm. This fills the gap of insufficient parameter optimization for deep multi-coal seam mining in existing studies. Up to the end of mining, the maximum deformation of the protected roadway is 8 cm, ensuring normal operation.

(3)

An innovative sealing scheme combining AB material grouting with high-strength detonator pins is developed, which shortens the sealing-to-blasting time to 10 min (far less than traditional cement sealing’s 24+ h) and improves sealing integrity. This solves the efficiency and reliability problems of existing deep-hole sealing processes.