Study on the mechanism and prevention techniques of dynamic disaster in nearly vertical extra-thick coal seams

Abstract

To address the issues of complex rock mass movement and dynamic disasters during the mining of near-vertical extra-thick coal seams, this study takes the + 425 level B3 + 6 working face of Wudong Coal Mine as the research background, aiming to investigate the mechanism of dynamic disasters. By adopting theoretical analysis, on-site investigation, and physical simulation, the study established a mechanical model for rock pillar deformation and failure, analyzed the failure modes of rock pillar as well as roof and floor, and their impacts on coal seam stability, revealed the corresponding disaster-inducing mechanism, and proposed a pressure relief and rock burst prevention technology. The results show that the deformation and prying rotation of rock pillar are the main factors causing strain energy accumulation in coal-rock masses, while the fracture of rock pillar and the toppling-sliding of roof and floor are the primary forms of dynamic impact loading. Specifically, the large-scale fracture depth of rock pillar reaches 350 m, and the fracture step distance of the immediate roof exceeds 43 m. To prevent and control rock burst, a pressure relief and rock burst prevention technology was proposed, which achieves pressure relief through deep-shallow hole blasting on the roof and floor. Combined with the numerical simulation analysis of deep-shallow hole blasting, the significant reduction in electromagnetic intensity, the gradual decrease in the frequency of microseismic events, and the gradual reduction in energy have verified the pressure relief effect of this technology. This study provides an effective technical approach for the safe mining of coal seams.

Introduction

“Nearly vertical extra-thick coal seam (NVEC)” refer to a specific type of coal seam with an inclination angle that approaches verticality. They are primarily distributed in western China, such as Xinjiang, Qinghai, Inner Mongolia, and other regions. Due to their near-vertical occurrence, the dip-direction component of rock gravity is significant. During sublevel mining, the stress distribution in the surrounding rock and the failure modes of the roof and floor differ fundamentally from those in near-horizontal or inclined coal seams. This leads to frequent dynamic disasters, severely constraining the safe and efficient extraction of coal resources in these regions. Compared with near-horizontal or inclined coal seams, steeply inclined seams exhibit more complex stress evolution and a higher susceptibility to dip-direction sliding and instability. Furthermore, their characteristic alternate mining process in coal seam groups intensifies the spatiotemporal superposition of mining-induced effects and increases the uncertainty surrounding disaster mechanisms. Consequently, existing research findings based on near-horizontal or single-seam mining are difficult to apply directly, highlighting an urgent need for separate systematic studies focusing on the unique dynamic disaster mechanisms and prevention technologies for such seams.

In recent years, domestic and international researchers have conducted extensive research on the dynamic disasters associated with the mining of NVEC and have made some progress in this field. The occurrence of dynamic disasters is directly related to the movement of overlying strata. In terms of overlying strata movement in nearly vertical and steeply inclined coal seams, researchers established a mechanical model for the breaking and tilting of the roof strata in steeply inclined coal seams. They found that during horizontal sublevel mining, two types of failure modes can occur: “toppling” in the upper sublevel and “slumping” in the lower sublevel¹. Researchers have studied the initial fracture characteristics of overlying strata in steeply inclined coal seams². Through the use of discrete element numerical modeling, researchers have examined the caving law of hard roofs in both horizontal and inclined coal seams³. The initial and periodic fracture phases of the roof are derived by developing a thin plate mechanical model of the roof stress distribution in the steeply inclined coal seam⁴. In the realm of dynamic disaster theory research, significant achievements have been made, with researchers proposing a multitude of theories including strength theory⁵, stiffness theory^6,7, energy theory^8,9, and bursting liability^10,11, among others. Building upon these theories, researchers have also made significant efforts to explore the interaction between dynamic and static load sources in the context of dynamic disaster prevention. For example, some researchers have suggested that when the sum of static and dynamic loads exceeds the minimum load required to induce rock burst, it can trigger dynamic pressure disasters^12,13. Researchers found that dynamic failure in steeply inclined thick coal seams primarily occurs near the roof. The clamping effect of the roof and floor, along with the rotation of the deep roof, leads to a higher risk of coal burst in the coal seam adjacent to the roof. After dynamic ground pressure occurs, fractures rapidly propagate into the deeper coal seam and the coal seam near the roof. Tensile fractures dominate in the shallow coal seam, while shear fractures dominate in the deeper coal seam¹⁴. The construction of a mechanical model for a steeply inclined thick coal seam allows for the analysis of the evolution of the stress and energy fields during mining. This model also helps determine the type of “dislocation constraint” dynamic disaster that occurs when the roof collapses and the working face is subjected to superimposed load¹⁵. Researchers established a model for overlying strata breaking and proposed that the source of dynamic load in steeply inclined thick coal seams is the instability of the roof-floor clamping structure¹⁶.

In the prevention and warning of dynamic disasters, researchers have observed that micro-seismic events monitored during the mining of steeply inclined extra-thick coal seams are primarily concentrated in the roof and rock pillars. They have proposed that the pressure and levering effects of rock pillars on the coal body are the primary causes of dynamic ground pressure incidents, and they have developed a multi-parameter early warning model for dynamic ground pressure^17–19. Researchers used the sharp increase or decrease in micro-seismic energy and event counts as precursor information for predicting dynamic ground pressure occurrences. They validated this warning method through acoustic emission experiments and demonstrated its effectiveness²⁰. Researchers discovered that the energy stored in steeply inclined rock masses is mainly released through low-energy micro-seismic events during deep mining, and the rock mass remains in a “low-frequency-energy-storage period” for an extended period. They proposed energy control measures, including “blasting-water injection” for sandwiched rock pillars and “deep-shallow” borehole blasting for the roof of the B3 + 6 coal seam²¹. Researchers introduced a method for pre-split blasting and unloading of rock pillars using deep hole blasting on the surface. This approach disrupts the integrity of the rock pillars to achieve the unloading effect²². Researchers utilized the discrete element numerical simulation method to study the collapse behavior of hard roof in both horizontal and inclined coal seams. Based on the characteristics of roof collapse, they designed pre-split blasting schemes for hard roof and achieved favorable results²³. For steeply inclined coal seam mining, Shengli Yang established the intrinsic relationship between coal wall spalling and various influencing factors, and proposed corresponding control measures for coal wall stability. This work provides theoretical and technical guidance for the prevention and control of coal wall spalling in steeply inclined coal seam mining²⁴.

In conclusion, a great deal of study has been done recently by researchers both domestically and internationally on the causes and prevention of dynamic disasters. The potential for disaster during the mining process is still quite significant due to the complexity of the dip angle and propensity in steeply inclined coal seam mining. The mechanism of dynamic disaster caused by alternating mining of two layers of the coal seam is more complex and variable, especially in the mining of horizontal sublevel fully mechanized caving face with NVEC. This suggests that more research is still needed to fully comprehend the mechanism of dynamic disaster and its prevention and control measures. Thus, the NVEC in Wudong Coal Mine’s south mining area provides the engineering background for this research, which then uses theoretical analysis, physical modeling, and numerical simulation to build a mechanical model of rock pillar deformation and collapse. This model reveals the mechanism causing the disaster in the case of horizontal sublevel fully mechanized caving mining by thoroughly analyzing the stress and migration law of the interlayer rock pillar, as well as the roof and floor of the working face. It also establishes the rock stratum migration model. A set of pressure relief and prevention of rock burst technology schemes for the NVEC are offered, based on the aforementioned study results. The plan is to lower the probability of dynamic disasters in the mining process and increase mining safety by offering references and guidelines for disaster prevention and management in the mining of NVEC under similar conditions.

Engineering background

Wudong Coal Mine is located in the eastern part of Urumqi City, Xinjiang, China. It mainly mines B1 + 2 and B3 + 6 coal seams. The average thickness of the B1 + 2 coal seam is 28 m, and the average thickness of the B3 + 6 coal seam is 40 m. These coal seams exhibit an inclination angle ranging from 85° to 87°, categorizing them as NVEC. Rock pillars separate the two coal seams, with the width of the rock pillars gradually narrowing from west to east, varying between 50 m and 110 m in thickness, as depicted in Fig. 1. The B1 + 2 and B3 + 6 working faces utilize the horizontal sublevel caving mining method. The B3 + 6 working face serves as the initial mining face, and the two coal seams are alternately exploited.

Fig. 1.

Section of the south mining area of the Wudong coal mine.

The southern mining area of Wudong Coal Mine is a typical NVEC mine. There have been three dynamic disasters caused by rock pillars in recent years, and one roof dynamic disaster has occurred, which has a serious impact on safe and efficient production, as shown in Table 1. Therefore, it is urgent to study the strata migration law of NVEC, especially rock pillars, and put forward the dynamic disaster prevention and control technology suitable for this condition.

Table 1.

Dynamic disaster incidents in the Southern mining area.

Mining level (m)	Hypocenter level (m)	Hypocenter location	Energy level (J)
+ 475	+ 480	Rock Pillar	5.0*108
+ 450	+ 460	Rock Pillar	9.5*106
+ 450	+ 460	B6 Roof	2.1*108
+ 450	+ 430	Rock Pillar	2.2*106

Deformation and failure of strata during the mining of NVEC

The dynamic disaster of NVEC is closely related to the movement of rock strata. In this section, the mechanical theoretical models are established for the hard rock pillar, B3 + 6 roof, and B1 + 2 floor of the NVEC, to describe the movement law of rock strata in the mining of NVEC.

Deformation and failure of rock pillars

Analysis of rock pillar deformation

A mechanical model for the deformation of rock pillars under the support of backfilled loess in the goaf of the NVEC was constructed using the southern extraction area of the Wudong Coal Mine as the engineering background, as shown in Fig. 2. The rock pillars, deprived of coal body support, were simplified into cantilever beams. The surface end of the rock pillar was regarded as the free end, and the point of confinement by the coal seam was set as the coordinate origin. The backfilled loess in the goaf of the B3 + 6 and B1 + 2 working faces was simplified as a spring-like support mechanism, providing support to the rock pillar²⁵. Under the influence of its weight, the rock pillar underwent bending deformation towards the B1 + 2 goaf.

Fig. 2.

Rock pillar mechanical model.

The total potential energy of the rock pillar, which includes the strain energy of the rock pillar, the elastic potential energy of the loose filler body in the goaf, and the work of the rock pillar load, based on the concept of strain energy and under the action of the roof and floor holding.

1

Where E represents the elastic modulus of the rock pillar, GPa; I stands for the moment of inertia of the rock pillar, with a value of bh³/12, b and h denote the width and average thickness of the rock pillar, respectively, (m); w represents the deflection curve of the rock pillar, (m); l signifies the distance from any point on the suspended rock pillar to the working face, (m); q represents the self-weight component per unit height of the rock pillar, MPa; k stands for the stiffness of the support spring in the goaf, kN/m; H denotes the depth of burial of the working face, (m); α represents the angle of inclination of the rock pillar, (°); γ₁ signifies the unit weight of the rock pillar, kN/m³.

According to the principle of minimum potential energy, the first variation of the total potential energy of the rock pillar structure at a stable state is 0.

2

The boundary condition of the rock pillar under the working face can be simplified as a fixed end. It can be known from the mechanics of materials that the displacement and rotation angle of the rock pillar is 0 under the fixed boundary condition. Substituting this condition into the previous equation, the deflection of the rock pillar can be solved as follows:

3

Where:

According to the field measured data, H = 375 m, q = γ₁hcosα, γ₁ = 24.83 kN/m³, h = 80 m, α = 85 °, E = 20 GPa, k = 1000 kN/m. The elastic modulus of the rock pillar, E = 20 GPa, was determined primarily through laboratory tests on rock specimens extracted from the rock pillar. The modulus was derived from the slope of the approximately linear portion of the recorded stress–strain curve. Based on a series of such tests, the elastic modulus for the corresponding stratum was established as 20 GPa. The support stiffness k = 1000 kN/m is an estimated value, as the in-situ backfill density is variable. In this analysis, the goaf backfilled with loess and gangue is idealized as a spring that resists deformation of the rock pillar, under the assumption of uniform backfill material. This stiffness value was obtained through the calculation procedure detailed in the preceding section. The physical and mechanical parameters of the above rock pillar are substituted into the equation to draw the curve of the offset of the rock pillar under different mining depth conditions. As shown in Fig. 3, when the mining depth is about 375 m, if the rock pillar is not broken, the rock pillar can be shifted to the B1 + 2 goaf by more than 0.5 m under its gravity.

Fig. 3.

Relationship between rock pillar offset and overhang distance.

Analysis of rock pillar failure

According to the deflection curve function of the rock pillar obtained by solving the mechanical model of thick and hard rock pillar supported by backfill loess, the stress distribution curve of the rock pillar can be obtained by solving the following equation. 

4

5

The final solution yields the stress distribution curve function of the rock pillar as follows:

6

Where:

Based on Brazilian splitting tests conducted on field cores, the tensile strength of the rock specimens was determined from the peak axial load. Statistical analysis of the data yielded an average tensile strength of 1.5 MPa, denoted as σ_b = 1.5 MPa. Given that the tensile strength of rock is significantly lower than its compressive strength, the primary failure mode of the rock specimens is considered to be tensile failure. When σ_x0 > σ_b, which means that when the deflection at the position x₀ exceeds the value of b, the rock pillar at that location undergoes failure.

Because the NVEC in Wudong mine mining southern area adopt horizontal sublevel fully mechanized caving mining, the sublevel height is 25 m. With the continuous mining of coal seam, the mining depth of the working face can be H=25i (i = 1, 2… 14), and the neutral axis distance y = h/2, and the stress distribution curve of rock pillar under this condition is drawn, as shown in Fig. 4. When i = 5, the maximum stress of the rock pillar does not exceed the tensile strength of the rock pillar, that is, the rock pillar has not been damaged. When i = 6, the maximum stress of the rock pillar is 1.88 MPa, which exceeds the tensile stress strength limit of the rock pillar. Therefore, when the coal seam with a buried depth of 150 m is mined, the right edge of the rock pillar begins to be damaged, but the damage range is not large, that is, the 150 m of the exposed length of the rock pillar is considered to be the initiation and failure point of the rock pillar.

Fig. 4.

Stress distribution of rock pillar at 150 m overhang distance.

It can be seen from Eq. (6) that with the increase of mining depth, the stress of each point of the rock pillar shows an increasing trend, and no matter how deep the mining depth is, the maximum stress distribution area of the rock pillar is located at the bottom of the rock pillar. Therefore, it can be found that there are three stress concentration zones under the clamping rock pillar structure of the B3 + 6 coal seam, which are the compressive stress area of the B1 + 2 coal seam, the prying stress zone of the B3 + 6 coal seam, and the stress concentration zone at the bottom of the rock pillar, as shown in Fig. 5.

Fig. 5.

Schematic diagram of rock pillar damage.

The stress distribution within the rock pillar under an unsupported span of 350 m is presented in Fig. 6. As the mining depth continues to increase, the hanging distance of the rock pillar becomes larger, leading to higher stress concentration at the base. Assuming the neutral axis distance y = h/10, when i = 14, the maximum stress at the base of the rock pillar reaches 2.05 MPa, exceeding the tensile strength limit. On the other hand, the maximum stress value at the location of the crack, 150 m away from the base of the rock pillar, is 0.68 MPa, which is significantly lower than the tensile strength limit. Therefore, when mining at a depth of 350 m, the extent of the failure section at the exposed 150 m of the rock pillar is not notably expanded. However, at the position h/10 near the neutral axis of the rock pillar, the base of the rock pillar experiences cracking. At this point, it can be deduced that the region below the initiation point of the crack in the rock pillar will experience varying degrees of tensile failure, particularly a substantial failure at the base of the rock pillar.

Fig. 6.

Stress distribution of rock pillar at 350 m overhang distance.

Analysis of rock pillar failure based on micro-seismic

The statistical analysis of micro-seismic distribution monitoring data in the hard rock pillars in recent years is shown in Fig. 7. When the mining depth reaches 350 m, around an elevation of approximately 450 m, there is a concentration of high-energy-level micro-seismic events within the rock pillars. Combining this with the analysis of the locations of pillar fractures derived in Sect. 3.1.2 and the occurrence of three rock pillar-induced disasters in recent years, it can be concluded that there is extensive internal damage within the rock pillars near the depth of 350 m. Ultimately, this leads to rock pillar fractures and triggers dynamic disasters. Furthermore, the analysis of micro-seismic data and dynamic disaster events to some extent validates the reasonableness and reliability of the rock pillar failure theory analysis.

Fig. 7.

Micro-seismic distribution monitoring results of rock pillars.

Analysis of roof and floor breakage

Mechanical analysis of roof breakage

Compared to other coal seams, a key difference in the case of NVEC lies in the different stress conditions of the roof. Due to the nearly vertical inclination of the roof in NVEC, the vertical components of the self-load of the roof and overlying rock load are very small. This suggests that the characteristics of roof breakage in NVEC should significantly differ from those in other coal seams. A mechanical model for the roof breakage of NVEC has been established to reveal the characteristics of roof breakage in such conditions²⁶. Figure 8 shows a mechanical model of a fixed-supported beam at both ends for the first section of roof breakage in NVEC.

Fig. 8.

Structural force analysis of fixed-supported beam at both ends. (a) Double-end rigid support beam structure, (b) Normal force, (c) Shear stress.

7

8

Where H represents the mining depth, m; l is the hanging distance, m; γ is the density of the roof, taken as 26.7 kN/m³; α is the angle of the roof, assumed as 87°; h is the thickness of the roof, taken as 5 m.

As shown in Fig. 9, the maximum tensile stress in the roof occurs at the upper surface near the top. Therefore, it can be assumed that the upper end of the roof will experience tensile failure first, while the lower end of the roof will break as the mining depth increases.

Fig. 9.

Roof stress analysis.

The maximum tensile stress value at the upper end of the roof is:

9

Setting σ_cmax ≥ 1.5 MPa, we can solve for the hanging step distance of the roof’s upper end when tensile failure occurs, which is 28 m.

After the upper end of the immediate roof breaks, it will bend and deform towards the goaf, generating separation from the basic roof. By simplifying the deforming suspended roof as a cantilever beam, it becomes evident that its mechanical model is approximately similar to the rock pillar deformation and failure mechanical models in Sect. 3.1.1 and 3.1.2.

Taking the immediate roof thickness as 5 m, dip angle as 87°, density as 26.7 kN/m³, and E as 20 GPa, substituting these parameters into Eq. (6) reveals that when the immediate roof is suspended for 43 m, as shown in Fig. 10, the tensile stress at the bottom of the roof exceeds 1.5 MPa, causing tensile failure. Therefore, the roof’s fracture step distance is 43 m. In reality, the upper end of the roof is not truly a free end but rather a simply supported end. To facilitate calculations, it was simplified as a free end. Therefore, the actual theoretical step distance for roof fracture should be greater than 43 m. Hence, it can be concluded that the initial fracture step distance of the immediate roof’s top in the southern mining area of the Wudong mine is greater than 43 m. From a mechanical theory perspective, this suggests the existence of a substantial suspended roof phenomenon in the mining of NVEC.

Fig. 10.

Relationship between the stress distribution of the roof and the overhanging distance.

As the depth of sublevel mining increases in the NVEC, factors such as coal seam dip angle, the height of fractured rock blocks, joint fissures, and others collectively influence the stability of the roof. Under these conditions, the roof may undergo either toppling or slumping forms of failure. Determining when toppling or slumping failures occur can be assessed using limit equilibrium analysis and the coefficient transfer method¹.

10

Where P_n−1,t and P_n−1,s represent the critical values of lateral force required to prevent tilting and sliding failure of the roof rock block n and the greater of the two is denoted as P_n−1. If P_n−1=P_n−1,t, it indicates that the roof rock block is prone to tilting failure, while if P_n−1=P_n−1,s, it indicates the potential for sliding failure of the roof rock block. If all roof rock blocks satisfy P_n−1=P_n−1,t, it implies that tilting failure will extend to the lowest roof rock block without sliding failure occurring. Conversely, if a certain roof rock block satisfies P_n−1=P_n−1,s, then all subsequent roof rock blocks, including that specific block, are in a critical state of sliding failure. According to Eq. (10), it is evident that a higher roof-to-thickness ratio increases the likelihood of toppling failure. Based on this, it can be inferred that roof failure in NVEC primarily occurs in the form of toppling, with the movement directed toward the goaf.

Similar simulation experiment of roof fracture

To investigate the roof fracture and movement patterns under different sublevel mining depths, this study conducted a similarity simulation experiment based on the geological setting of the nearly vertical ultra-thick coal seam group in the southern mining area of Wudong Coal Mine. Guided by similarity theory, the geometric dimensions, similarity ratios, and boundary conditions of the experimental model were determined. A two-dimensional test platform (length × width × height = 180 cm × 16 cm × 140 cm) was employed, with a geometric similarity ratio of 1:150 and a unit weight similarity ratio of 1:1.7. The dip angle of the strata was set to 85°. Due to the nearly vertical orientation of the coal seams, horizontal sublevel mining was adopted, with 30 cm-wide boundary coal pillars reserved on both sides. The model was constructed to a height of 114 cm, corresponding to a real-world height of 171 m. A total of six sublevels were mined, each with a height equivalent to 25 m in the field. To account for boundary effects, a 21 m-thick coal seam (in prototype scale) was left at the bottom of the model. Overburden pressure was simulated by applying counterweights on top of the model to replicate realistic in-situ loading conditions, as illustrated in Fig. 11. Based on the similarity ratios and model dimensions²⁷, the material proportions for the similarity simulation experiment were determined, as summarized in Table 2.

Fig. 11.

Similar simulation test model.

Table 2.

Similar simulation test model material.

Layer number	Lithology	Thickness/m	Simulation thickness (cm)	Ratio	Sand (Kg)	Lime (Kg)	Gypsum (Kg)
1	Boundary	10	5.0	655	8.067	0.672	0.672
2	Muddy sandstone	21	10.5	755	8.092	0.578	0.578
3	Silt sandstone	9	4.5	655	5.378	0.448	0.448
4	B1 + 2 coal	30	15.0	855	4.815	0.301	0.301
5	Grey mudstone	4	2.0	764	5.267	0.451	0.301
6	Silt sandstone	11	5.5	655	5.378	0.448	0.448
7	Muddy sandstone	16	8.0	755	5.394	0.385	0.385
8	Silt sandstone	19	9.5	655	5.378	0.448	0.448
9	Muddy sandstone	19	9.5	755	5.394	0.385	0.385
10	Silt sandstone	11	5.5	655	5.378	0.448	0.448
11	B3 + 6 coal	50	25.0	855	4.815	0.301	0.301
12	Grey mudstone	5	2.5	764	5.267	0.451	0.301
13	Silt sandstone	25	12.5	655	8.067	0.672	0.672
14	Boundary	10	5.0	655	8.067	0.672	0.672

In this experiment, a total of six sublevels were excavated to simulate a mining depth of 150 m. The B3 + 6 and B1 + 2 working faces were alternately mined, with B3 + 6 being the initial mining face. During the first sublevel excavation, there were no apparent deformation or failure observed in the roof strata of both B3 + 6 and B1 + 2 working faces. It was only in the second sublevel excavation of the B1 + 2 working face that a sudden rotation of the overhanging roof strata of the B3 + 6 coal seam occurred along the bottom support point of the rock block. This eventually led to tilting, folding, and breakage of the strata, resulting in the rock block collapsing entirely onto the top coal layer. The initial failure occurred at distance of approximately 50 m. This phenomenon is depicted in Fig. 12. Consequently, the failure mode of the roof strata in the upper sublevel mainly manifested as toppling.

Fig. 12.

Second sublevel mining. (a) In the second sublevel of B3 + 6, (b) In the second sublevel of B1 + 2.

When reaching the third sublevel excavation of the B3 + 6 working face, the collapsed strata from the upper sublevel as well as the remaining top coal move downward into the goaf, as shown in Fig. 13. The collapsed coal and rock blocks, after falling and rotating, undergo further fragmentation, resulting in a reduction in their height. Upon reaching the third sublevel excavation of the B1 + 2 working face, the overlying strata of the B1 + 2 coal seam directly above undergo tilting and folding, breaking off. The periodic fracture step distance remains at 50 m, the same as the initial fracture step distance. However, there still exists an overhanging roof in the lower part of the second sublevel.

Fig. 13.

Third sublevel mining.

Upon reaching the fourth sublevel excavation, the immediate roof of the B3 + 6 goaf experiences fracture and collapse, while the immediate roof of the B1 + 2 coal seam undergoes delamination. The collapsed roof strata from above and the remaining top coal experience a secondary failure during their downward movement, resulting in a two-stage fracture and accumulation above the top coal.

As shown in Fig. 14, when advancing to the fifth sublevel of the B1 + 2 coal seam, the collapsed roof strata from the goaf and the remaining top coal continue to move downward. The overhanging roof strata of the B1 + 2 coal seam undergo fracture and tilting, forming a triangular support structure with the bottom of the B1 + 2 coal seam. This supports the collapsed and fragmented rock layers above, leading to widespread delamination of the overlying immediate roof strata. Simultaneously, the overhanging roof strata of the B3 + 6 coal seam also experience fracture. Due to the greater length of the B3 + 6 working face and the larger movement space of the rock layers, the fractured rock layers tilt and break into blocks, which accumulate above the goaf. However, no distinct structural pattern is formed.

Fig. 14.

Fifth sublevel mining.

Upon reaching the sixth sublevel excavation of the B1 + 2 coal seam, the overhanging roof strata with a stepped profile above the B3 + 6 goaf toppls and breaks, forming fractured rock blocks that spread out over the goaf. As the working face continues to advance, a substantial collapse occurs in the immediate roof of the B1 + 2 coal seam above the goaf. This leads to the formation of a triangular support structure in the upper section, which undergoes instability due to the impact and experiences further fracturing during its downward movement. Ultimately, this results in a collapsed structure as shown in Fig. 15.

Fig. 15.

Sixth sublevel mining.

Additionally, due to the relatively short length of the B1 + 2 working face, when a large-scale collapse of rock layers occurs, the movement space is limited. In this scenario, the roof strata tilt towards the goaf and then come into contact with the rock layers of the working face’s floor. Subsequently, the rock blocks near the floor undergo fracture. These rock blocks may slide along the floor or fall vertically downward.

In summary, the rock breaking of horizontal sublevel mining in NVEC is mainly manifested as the separation of the immediate roof strata under the influence of mining, toppling under the action of gravity, and the broken blocks filling the goaf. With the increase of sublevel mining depth, the immediate roof strata gradually fall layer by layer, and the residual top coal and gangue mixed with rock blocks in the goaf migrate downward, mainly in the form of toppling, falling, and sliding, as shown in Fig. 16. Among them, the rock blocks near the roof of the working face are mostly toppling or sliding downward. The rock blocks near the floor of the working face are mostly sliding downward, and the rock blocks in the middle of the goaf are mostly falling and turning. It will have a great impact on the lower segment coal body and working face.

Fig. 16.

Mining process roof collapse “most dangerous state”.

Analysis of the floor breakage

As shown in Fig. 17, a mechanical model for the immediate floor slide of the NVEC has been established. Gsinα represents the gravitational component of the upper floor; F_N1 and F_N2 are the support forces exerted by the immediate floor from the main floor of the empty section and the coal compression section of the working face respectively; f represents the frictional force between the main floor and the immediate floor; F_A denotes the support force on the lower end of the floor; q(x) represents the compressive force exerted by the coal on the immediate floor.

Fig. 17.

Mechanical analysis model of the floor.

When the mining depth of the B1 + 2 working face reaches a certain level, there will be a significant concentration of stress on the floor below the working face. The primary source of this stress concentration is the force q(x) transmitted through the coal seam, causing rock strata to compress, as well as the effect of its weight. The horizontal stress concentration beneath the working face floor increases with the increasing mining depth of the working face, and the peak stress concentration also increases accordingly. In the section of the floor strata beneath the working face that experiences coal seam compression, the rock strata will undergo plastic yielding under high-stress conditions. New cracks will form within the rock strata, and existing cracks will also expand. When the floor strata under the working face enter the goaf stage, they suddenly experience unloading of the high load they previously bore. This sudden unloading can lead to significant rebound deformation, and a substantial decrease in the strength of the floor strata, and this phenomenon becomes more severe as the mining depth increases. When the strength of the floor strata decreases to a level where overall failure is likely, it becomes susceptible to various factors, including the roof collapse of the working face or significant movement of rock pillars. Under these conditions, the floor beneath the working face will experience shear failure due to its weight, leading to overall collapse along shear planes. Furthermore, the periodic rupture of rock pillars alters the leveraging and squeezing effects of these pillars on the coal mass, consequently impacting the form and extent of cyclic failures in the floor. Investigating the periodic floor failures will be a focal point in the subsequent research.

The preceding analysis has examined the deformation and failure characteristics of hard rock pillars and the roof and floor in the mining of NVEC from a mechanical perspective. Based on this, a rock strata movement model is established for the mining conditions of NVEC, as illustrated in Fig. 18. Specifically, in the conditions of mining NVEC, when the roof of the B3 + 6 coal seam first undergoes sublevel rupture, it primarily experiences toppling failure. During sublevel rupture, there is a high probability of toppling failure, but slumping failure may also occur, depending on factors such as coal seam dip angle and joint fractures. In contrast, the floor of the B1 + 2 coal seam may experience slumping failure when subjected to the combined effects of gravity and lateral coal pressure during both initial and subsequent sublevel ruptures. The hard rock pillars continuously deform and displace towards the goaf area of the B1 + 2 coal seam and ultimately undergo extensive failure at the base of the rock pillars.

Fig. 18.

Strata movement model of NVEC mining.

Analysis of dynamic disasters

Based on the above overburden deformation and failure research, the disaster-causing force source of NVEC mining is analyzed.

Analysis of static loading sources

As the established mechanical model of rock pillar deformation, it is evident that rock pillars within the NVEC will undergo bending deformation towards the B1 + 2 goaf, exerting prying, twisting, and compressive effects on the coal masses on both sides. Referencing Fig. 19, it is observed that with the increase in mining depth, the displacement of the rock pillars gradually increases, leading to a progressive intensification of prying and compressive effects on the unmined coal seam area.

Fig. 19.

Schematic diagram of the coal body being pried and squeezed by the rock pillar.

To quantify the prying and squeezing action of the rock pillar on the coal bodies on both sides, it is approximately assumed that the displacement and deformation of the coal seam beneath the B3 + 6 working face exhibit a proportional relationship with the deformation of the exposed rock pillar, as illustrated in Fig. 20. Through analysis, the relationship is determined as follows:

11

Fig. 20.

Proportional deformation relationship between coal seam and rock pillar.

To mitigate the errors introduced by the idealized model, when the burial depth is H, consider , and perform a mechanical analysis on the rock pillar at point C located 30 m below the B3 + 6 working face.

12

13

Where l_b represents the length between point C on the rock pillar and the prying point, w_b represents the displacement of the rock pillar at point C, and h_B3+6 represents the average thickness of the B3 + 6 coal seam.

By combining the equations, an approximate derivation of the stress equation for compressive stress concerning the overhang distance L of the rock pillar can be obtained:

14

By substituting the physical and mechanical parameters of the rock pillar and coal seam into the above equation, the distribution curve of compressive stress concerning the overhang distance of the rock pillar can be obtained, as shown in Fig. 21. With the increase in mining depth, the deformation of the rock pillar increases, leading to a gradual increase in the compressive stress applied to the coal seam. The concentration of horizontal stress within the coal seam also rises, causing the accumulation of strain energy within the coal mass. Similarly, the immediate roof of the B3 + 6 coal seam has a similar effect on the coal mass. Ultimately, this lays a strong static load foundation for the occurrence of dynamic disasters.

Fig. 21.

Relationship between extrusion stress of coal body by rock pillar and overhang distance of rock pillar.

Analysis of dynamic loading source

As mentioned in Sect. 3.1.2 and 3.2, the failure of rock pillars and the collapse of the roof are the main dynamic load sources in the mining of NVEC. When the mining depth of the working face reaches 350 m, the rock pillars in the NVEC experience large-scale failure. This leads to the abrupt release of a significant amount of stored strain energy, and the released elastic strain energy is approximately equal to the stored bending strain energy within these rock pillars. According to the following equation, the more strain energy is released, the greater the impact load, resulting in a stronger dynamic effect on the coal seam:

15

Considering the cushioning effect of fragmented top coal and gangue on impact energy, the impact load induced by the large-span suspended roof can be determined by Eq. (16). It can be inferred from this equation that the impact load exhibits a positive correlation with roof height, roof thickness, and impact height.

16

Furthermore, with the continuous increase in mining depth, the influence of compressive-shear plastic yielding and unloading on the floor becomes more significant. Consequently, the risk of floor failure increases. Therefore, the floor represents a potential dynamic load source for dynamic disasters.

In conclusion, the occurrence of dynamic disasters in the NVEC can manifest in various combinations of circumstances. However, the underlying principles remain consistent: the movement of rock pillars and the roof serves as the primary source for dynamic hazard occurrence. The mechanism underlying these hazards is based on the deformation and prying of rock pillars and roof, which provide a strong static loading foundation for dynamic hazard occurrence. The dynamic perturbation from the rupture of rock pillars and roof leads to internal stress within the coal mass reaching or exceeding the critical stress threshold for dynamic hazard manifestation, ultimately causing dynamic hazard occurrence. Therefore, the roof and rock pillars are the primary focus of current research on dynamic hazard prevention and control. Additionally, the floor, as a potential hazard factor, requires increased research attention.

Pressure relief and prevention rock burst technology

Through the above research, we understand that the occurrence of dynamic disasters is the result of the combined effects of high static stress concentration and dynamic impact disturbance. If we can propose a technology to mitigate the concentration of horizontal stress by dynamic impact, it may be possible to prevent dynamic disasters from occurring. Based on this approach and considering the characteristics of dynamic and static load sources in the NVEC, we have specifically proposed the roof and floor deep-hole blasting unloading technology as a targeted measure to prevent dynamic disasters^28–31.

Deep and shallow hole blasting pressure relief scheme

As shown in Fig. 22, deep and shallow hole blasting is conducted in the roof and floor on both sides of the working face, creating a three-dimensional “buffer zone” in the roof and floor. This alters the propagation path of horizontal stress, obstructing and attenuating stress transmission, thus reducing the effect of the maximum horizontal principal stress. It redirects the influence of the maximum horizontal principal stress away from the mining activity area. Additionally, when internal damage occurs within the rock pillar, the “buffer zone” in the roof and floor also hinders the release of energy or vibration transmission from the rock pillar, preventing the completion of the dynamic disaster process.

Fig. 22.

Principle of roof and floor pressure relief blasting in NVEC.

In the south mining area of Wudong Coal Mine, a deep and shallow hole blasting pressure relief scheme for the roof and floor is implemented. The scheme is based on the research achievements of ground pressure control theory in the context of NVEC and is tailored to specific site conditions. The detailed scheme is as follows:

For shallow layer blasting holes, each group consists of 3 boreholes with designed lengths of 25 m, 25 m, and 35 m. Borehole angles are set at 25°, 45°, and 60°, respectively. Explosive charge lengths are 15 m, 15 m, and 23 m, with a borehole diameter of 113 mm. In moderate impact hazard zones, the borehole spacing is 7.5 m per row. In weak impact hazard zones, the borehole spacing is 10 m per row.

For deep-layer blasting holes, each group includes 2 boreholes with a designed length of 50 m and a borehole diameter of 113 mm. The explosive charge length is 20 m, and the stemming length is 30 m. Borehole angles during drilling are 25° and 35°, respectively. In moderate impact hazard zones, the borehole spacing is 22.5 m per row. In weak impact hazard zones, the borehole spacing is 30 m per row.

The blasting pressure relief is conducted at approximately 60 m ahead of the working face, and the arrangement of deep and shallow holes in the roof and floor is depicted in Fig. 23.

Fig. 23.

Schematic diagram of deep and shallow hole arrangement.

Analysis of the effect of deep and shallow hole blasting

To investigate the influence of deep and shallow hole blasting on the stress distribution within the coal mass and surrounding rock of the working face, this study established a numerical model for deep and shallow hole blasting under the mining conditions of the nearly vertical coal seams at Wudong Mine, based on FLAC3D, as shown in Fig. 24. The model dimensions were set to 400 m in length, 400 m in width, and 300 m in height, with the Mohr–Coulomb constitutive model adopted for the entire model. In terms of boundary conditions, roller supports were applied to the four lateral boundaries, while the bottom boundary was fixed, simulating the mechanical constraints of the in-situ rock mass. A vertical load of 5 MPa was applied to the upper surface to equivalently represent the gravitational stress of the overburden from the + 600 level to the ground surface³².

Fig. 24.

Numerical model for deep and shallow hole blasting.

The simulation took the surrounding rock state after the completion of mining at the two working faces of the + 450 level as the initial condition, and then progressively simulated the stepwise extraction of the B3 + 6 working face at the + 425 level. The parameters for deep and shallow hole blasting were designed according to the actual field practices at Wudong Mine. The blasting-influenced zones in the roof and floor were determined based on the layout angle, charge length, and blasting pressure relief radius, with the specific extent illustrated in Fig. 25.

Fig. 25.

Blasting simulation layout. (a) Deep hole arrangement, (b) Shallow hole arrangement.

Based on the field blasting parameters and the blasting pressure relief radius, the influenced zones on both the roof and floor sides of the working face were delineated. Under the spatial condition that blasting precedes the working face by 40 m, during the stepwise mining process, the mechanical parameters of the unit cells within the blasting-influenced zones of the roof and floor, located within 40 m ahead of the working face, were simultaneously reduced. This involved decreasing their cohesion and internal friction angle to simulate the damage-induced weakening of the coal and rock mass after blasting, thereby forming a “stress buffer zone”. The simulation employed a coupled “stepwise mining-stepwise weakening” approach to replicate the field construction sequence of simultaneous mining and blasting for pressure relief. The specific scheme is detailed in Table 3.

Table 3.

Numerical simulation of deep and shallow hole blasting.

Plan	The simulation content	Spacing	Reduction coefficient
1	No blasting	/	1
2	Deep hole blasting	10 m	0.6
3	Shallow hole blasting	10 m	0.6
4	Shallow and deep hole blasting	30 m; 10 m	0.6

To validate the blasting unloading theory, taking the example of advancing the working face to 100 m, stress data from the + 425 horizontal sublevel was extracted using Fish language. Then, using Origin software, three-dimensional stress distribution plots of the surrounding rock under four different simulation scenarios were generated for both working faces.

The horizontal stress distribution of the surrounding rock in the + 425 level working face under different simulation scenarios is shown in Fig. 26. Compared to the non-blasting scenario, blasting causes a certain range of unloading zones to appear on both sides of the working face. By conducting advanced blasting in the two tunnels in front of the working face, the horizontal stress inside the coal seam significantly decreases. The horizontal stress in the unloading zones is transferred to the non-blasting areas, forming two “protrusions” in the roof and floor, which are stress concentration zones resulting from the redistribution of stress after unloading.

Fig. 26.

Horizontal stress distribution in the surrounding rock under different scenarios. (a) No blasting, (b) Deep hole blasting, (c) Shallow hole blasting, (d) Shallow and deep hole blasting.

To more intuitively analyze the pressure relief effect of roof and floor blasting on the coal mass at the working face, a monitoring line was arranged along the center of the working face. Horizontal abutment pressure distribution curves under four working conditions, namely no blasting, deep-hole blasting, shallow-hole blasting, and combined deep- and shallow-hole blasting, were plotted, as shown in Fig. 27. Pressure relief blasting effectively reduces the horizontal abutment pressure in the coal mass ahead of the working face. Taking the horizontal stress approximately 100 m ahead of the working face as an example, the stress at this location under the no-blasting condition is about 5.0 MPa. After deep-hole blasting, the stress decreases to about 4.5 MPa, a reduction of approximately 10%. Following combined deep- and shallow-hole blasting, the stress further decreases to about 4.2 MPa, a reduction of approximately 16%. After shallow-hole blasting, the stress drops to about 4.0 MPa, representing a reduction of about 20%. A further comparison of the pressure relief effects under different blasting methods indicates that shallow-hole blasting outperforms combined deep- and shallow-hole blasting, which in turn is more effective than deep-hole blasting alone.

Fig. 27.

Horizontal stress in the coal body in front of the workings under different blasting schemes.

Field monitoring and data analysis

Electromagnetic radiation test

Electromagnetic radiation monitoring technology is an important tool for the early warning and prevention of coal mine dynamic disasters such as rock burst and coal and gas outbursts. Based on the response relationship between electromagnetic radiation signals generated during coal-rock deformation and fracturing and the stress state, researchers have established corresponding prediction principles and methods. A series of electromagnetic radiation monitoring instruments have also been developed, promoting the widespread application of this technology in mine sites^33–36.

Theoretical and experimental studies have shown that the intensity of electromagnetic radiation signals exhibits a significant positive correlation with the stress concentration level in the coal-rock mass. Therefore, changes in stress concentration can be indirectly quantified by measuring variations in electromagnetic radiation intensity before and after blasting for pressure relief. A decrease in electromagnetic radiation intensity after blasting indicates a reduction in stress concentration within the coal-rock mass, demonstrating the effectiveness of the pressure relief measures. Conversely, no significant change suggests limited effectiveness of the pressure relief operation. To evaluate the actual effectiveness of the blasting pressure relief technique adopted in this study, in-situ electromagnetic radiation signal tests were conducted on both sides of the roadway at the B3 + 6 working face before and after blasting. The tests employed a YDC7.4 portable electromagnetic radiation monitor for coal-rock dynamic disasters, as shown in Fig. 28. Monitoring points were arranged at 10-meter intervals along the roadway, and electromagnetic radiation data were collected both before and after blasting. The test results are presented in Fig. 29.

Fig. 28.

The YDC7.4 portable coal-rock dynamic disaster electromagnetic radiation instrument.

Fig. 29.

The electromagnetic intensity of the coal-rock mass before and after blasting.

Based on the changes in the electromagnetic intensity curves of the coal-rock mass before and after blasting, it can be observed that within the range of 1110 m–1087.5 m, after implementing shallow-hole and deep-hole blasting, there is a significant decrease in electromagnetic radiation intensity for both coal and rock on both sides of the B3 roadway. In contrast, in the area ahead of 1087.5 m where no blasting occurred, there is no significant difference in electromagnetic intensity between the two measurements. To analyze the changes in electromagnetic intensity before and after drilling and blasting within the range of 1110 m–1087.5 m for comparison, data from three measurement points, namely 1110 m, 1100 m, and 1090 m, within the blasting area were selected for analysis. The results show that after blasting, the electromagnetic intensity within the rock mass decreased by 29.5%, while the electromagnetic intensity within the coal mass decreased by 12.8%. The decrease in electromagnetic intensity in the coal-rock mass indirectly reflects a reduction in stress concentration within the coal-rock mass. Shallow-hole and deep-hole blasting had a good unloading effect on the coal-rock mass on both sides of the roadway, with the effect being more pronounced in the rock mass compared to the coal mass.

Analysis of micro-seismic monitoring data

Currently, mine rock burst monitoring and early warning systems make extensive use of the micro-seismic monitoring system. In Wudong Coal Mine’s southern mining region, the B3 + 6 working face is where the ARAMIS M/E micro-seismic monitoring system is located. The daily micro-seismic frequency and energy variations may be derived as indicated in Fig. 30 by selecting the micro-seismic data from about a month before and following the working face blasting. The temporal pattern of micro-seismic frequency and energy is complex. Based on a data study of both shallow and deep hole blasting, the working face’s micro-seismic energy and frequency rise before and following blasting, before decreasing. The process of deep and shallow hole blasting increases the frequency and energy of micro-seismic occurrences in the working face by widening the fissures in the surrounding rock, weakening its integrity, and releasing built-up stress. Following the release of tension, the surrounding rock tends to become more stable, micro-seismic event frequency gradually drops, and energy progressively declines. These observations show that deep and shallow hole blasting has a favorable pressure relief impact on the working face.

Fig. 30.

Variation of micro-seismic frequency and energy on site.

Conclusions

This study elucidates the mechanism and prevention principles of dynamic disasters during fully mechanized top-coal caving mining in steeply inclined, ultra-thick coal seam groups. Under mining disturbance, thick and hard rock pillars and the roof undergo deformation first, accumulating substantial elastic energy. When these structures fail, the accumulated energy is released abruptly and transferred to the coal mass at the working face, constituting the primary dynamic load for disasters such as rock burst. Accordingly, the deep and shallow hole blasting technique was proposed. Its fundamental principle lies in proactively creating pressure-relief zones and weakened bands to block or mitigate this “energy accumulation – release – transfer” pathway, thereby achieving active prevention and control of such disasters. Based on this, the main conclusions of this study are as follows:

1. We constructed a comprehensive mechanical model to investigate the deformation and failure of thick and hard rock pillars in NVEC. This model unveiled the mechanism of rock pillars exerting a squeezing and levering effect on the coal body. The model identified an initial failure depth for the rock pillars at 150 m and a large-scale failure depth at 350 m. Additionally, when combined with the analysis of micro-seismic data, these findings to some extent corroborated the validity and reliability of the rock pillar failure theory.

2. A mechanical model for roof and floor breakage in the vicinity of the NVEC has been established. The possibility of floor slumping under the action of pressure-shear-unloading cycles was determined. Combined with similar simulation experiments, the ultimate breaking step of the immediate roof was found to be greater than 43 m, and the phenomenon of a large-height overhanging roof in the NVEC was discovered. Toppling destruction was identified as the primary fracture motion mode of the roof. Finally, a strata movement model for the mining of the NVEC was proposed.

3. The bending deformation of rock pillars serves as the primary static load source for dynamic disasters in the working face. Large-scale destruction of rock pillars and high-level overhanging and tilting fractures in the roof are the main dynamic factors contributing to the formation of dynamic disasters. Meanwhile, floor collapse is a potential dynamic hazard factor that may occur, revealing the mechanism behind dynamic disasters in the conditions of horizontal sublevel top coal caving mining in NVEC.

4. The NVEC roof and floor blasting unloading and prevention of rock burst technique, proposed here, has been introduced to hinder and diminish the transfer of horizontal stress and the release of energy from far-field rock pillars to the coal body in the working face. Through numerical simulation analysis and on-site monitoring, it was discovered that the deep and shallow hole blasting technique effectively mitigated the concentration of horizontal stress in the coal-rock mass on both sides of the roadway. This, in turn, contributed to significant unloading effects within the areas adjacent to the working face and the roadway, as well as the good prevention of dynamic disasters.

Author contributions

Y. Z.: Methodology, Project administration, Resources. Q. L.: Methodology, Software, Writing—original draft. LH. L.: Supervision, Validation, Funding acquisition. J. Z.: Methodology, Writing—original draft. CY. L: Methodology, Software. H. Y: Supervision.

Funding

This work was supported by the China National Key R & D Program (2022YFC2904001) and the Fundamental Research Funds for the Central Universities (No. 2023YQTD02).

Data availability

The data presented in this study are available on request from the Corresponding author.

Declarations

Competing interests

The authors declare no competing interests.