Front Abutment Pressure Research Articles

Boundary quantitative characterization of the top-coal limit equilibrium zone in fully mechanized top-coal caving stope along the strike direction of working face

In the process of fully mechanized top-coal caving mining, the top-coal is affected by mining-induced stress, and the stress varies along the strike direction of working face, so the boundary position of its entering the limit equilibrium state changes accordingly. The determination of the boundary along the strike direction of working face can provide scientific guidance for the stability control of support-surrounding rock in fully mechanized top-coal caving face. Using the research methods of theoretical analysis, physical similarity simulation experiment and numerical simulation experiment, the stress state analysis model of the boundary position of the top-coal limit equilibrium zone under macro-scale conditions was established, the stress state characterization method of the boundary of the top-coal limit equilibrium zone along the strike direction of working face was given, and the quantitative characterization of the boundary of the top-coal limit equilibrium zone along the strike direction of working face was realized by combining with the mining-induced stress path, and the distance relationship between the boundary of the top-coal limit equilibrium zone and the langwall face along the strike direction of working face was revealed. The results show that after critical mining in fully mechanized top-coal caving face, the distance between the boundary of top-coal limit equilibrium zone and the langwall face along the strike direction of working face presents a relationship of increasing from top to bottom. The distance between the top-coal upper boundary and the langwall face was 2.85 m and the distance between the top-coal lower boundary and the langwall face was 5.39 m. The boundary of top-coal limit equilibrium zone along the strike direction of working face was verified by the top-coal elastic–plastic zone boundary and the boundary of the peak position of front abutment pressure in different layers of top-coal. The results show that the quantitative characterization of the top-coal limit equilibrium zone boundary along the strike direction of working face was reasonable. In order to improve mine production efficiency, optimization measures were put forward for hard coal seam and soft coal seam respectively.

Study on Characteristics of Front Abutment Pressure and Rational Stop-Mining Coal Pillar Width in Large Height Working Face

Study on Characteristics of Front Abutment Pressure and Rational Stop-Mining Coal Pillar Width in Large Height Working Face

Investigating the Mechanism of Strong Roof Weighting and Support Resistance Near Main Withdrawal Roadway in Large-Height Mining Face

Abstract Aiming at investigating the strong roof weighting when the large height mining face is nearing the main withdrawal roadway, the 52,304 working face (WF) nearly through the main withdrawal roadway mining in a colliery of Shendong coalfield was taken as the research background. The ground pressure, roof structure, and superposition effect of stress in the last mining stage were studied by field measurement, physical simulation, and numerical calculations. The obtained results demonstrated that the main roof formed the “long step voussoir beam” structure under the influence of the main withdrawal roadway. The superposition effect of the front abutment pressure of the WF and the concentrated stress of the main withdrawal roadway caused the stress asymmetrical distribution on the two sides -level hard rock straof the main withdrawal roadway, and the stability of the pillar on the mining side decreases. The initial average periodic weighting interval was 20.7 m. While the WF approaches the main withdrawal roadway, the pillar near the WF of the main withdrawal roadway collapsed, the main roof was broken ahead of the WF, and the actual roof control distance of support and the periodic weighting interval increased by 2.56 and 1.26 times the normal state, respectively. Consequently, the “static load” of the immediate roof and the “dynamic load” of the sliding unsteadiness of the long step voussoir beam increased. The structural model of the “long step voussoir beam” under the superposition of “static and dynamic load” was established concerning those results, and an expression was proposed to compute the support resistance. Meanwhile, the mechanism of strong roof weighting was revealed when the WF was nearly through the main withdrawal roadway. The research conclusion is expected to provide a guideline for the safe withdrawal of the large-height mining faces under similar conditions.

Simulation Study and Engineering Application of Weakening Mine Pressure Behavior in Stope through Ground Fracturing Thick and Hard Rock Strata

Fracturing hard roofs by ground hydraulic action is an important control technology for the strong mine pressure in the stope. In this paper, a new simulation method, “separate + interface,” is proposed, and two physical simulation experiments are conducted; the phenomenon of increased goaf pressure and decreased front abutment pressure is discovered after fracturing in the key strata, and then the influence of different fractured crack shapes on the front abutment pressure and the goaf stress is revealed. The results are as follows: Firstly, the separation under the high-level hard strata blocks the transmission of overburden load to the goaf, leading to the high-stress concentration of the coal seam, which is the main reason for the large deformation of roadways and the breakage of a single hydraulic prop in the roadway. Secondly, the weakening effect of mine pressure differs when hard rock strata are fractured artificially with different types of cracks. The peak value of abutment pressure is reduced from 24.91 to 20.60 MPa, 17.80 MPa, and 16.13 MPa with the vertical crack spacing of 20 m, 15 m, and 10 m, respectively, and the related goaf pressure is increased from 2.61 to 3.54 MPa, 3.91 MPa, and 4.34 MPa, respectively. The peak value of abutment pressure decreased from 24.79 to 22.08 MPa, 19.88 MPa, and 17.73 MPa. The related goaf pressure increased from 2.61 to 3.39 MPa, 3.81 MPa, and 4.43 MPa, respectively, with the key strata also fractured into two horizontal layers, three horizontal layers, and four horizontal layers with horizontal fractures. Thirdly, after the hard roof is fractured above the No. 8202 working face, the first breaking step distance of the main roof decreased from 112.6 to 90.32 cm, while the first breaking step distances of KS2 and KS3 decreased from 106.3 and 135.8 cm to 93.5 cm and 104.8 cm, respectively, and the goaf pressure also increased. Compared to the adjacent unfractured No. 8203 working face, the mine pressure intensity of the No. 8202 working face is significantly reduced. The research results can provide useful guidance for the treatment of strong mine pressure.

Distribution Law of Abutment Pressure and Support Technology in the Roadway of Repeated Mining in the Fully Mechanized Top Coal Caving Mining Face

Due to repeated mining in the 130203 fully mechanized top coal caving mining face of the coal mine, the severe mining pressure phenomenon occurs in the return airway. In order to determine the stress distribution state of the repeated mining roadway and address the issue of significant deformation in the surrounding rock, we utilized fast lagrangian analysis of continua in 3D (FLAC3D) software along with a two‐dimensional simplified model of advance abutment pressure. The study proposes the combined support forms of unit support, sectionable remote pumping prop, and hydraulic single prop. Meanwhile, the supportive effect of the advance return airway is being monitored. The results show that in the initial stage of mining face, the peak value of advancing abutment pressure on the fully mechanized caving face increases continuously, and the distance between the location of the peak stress and the coal wall becomes larger and larger. When the mining face reaches 60 m, the peak value of advancing abutment pressure gradually stabilizes, and the simulated peak value of advancing abutment pressure is 47.8 MPa. It is concluded that the area of sharp increase and decrease of the advancing abutment pressure is within 0–20 m from the coal wall, the distance between the peak abutment pressure and the coal wall is 10.8 m, and the strengthening supporting distance of the roadway affected by the front abutment pressure is 120 m. During the monitoring period, it was determined that the rate of movement of the roof and floor of the roadway under the combined support was 16.65 mm per day. This combined support method has a stable controlling effect on the surrounding rock of the roadway.

The Active Roof Supporting Technique of a Double-Layer Flexible and Thick Anchorage for Deep Withdrawal Roadway under Strong Mining Disturbance

Due to the strong disturbance of a mining face, the surrounding rock of the withdrawal roadway is susceptible to deformation and failure, which restricts the safe and efficient evacuation of mining equipment. To resolve this longstanding technical problem in mine production, an engineering investigation, numerical simulation, theoretical analysis, and other research methodologies were conducted in this study. Furthermore, the influence mechanism of mining-induced stress on the withdrawal roadway was revealed, the anti-disturbance principles of thick-layer anchorage of roadway roofs were elucidated, and a novel double-layer flexible support technique was proposed. The front abutment pressure, stress superposition, damage accumulation of the surrounding rock, and the fluctuation of mining-induced stress are the primary factors contributing to the significant deformation of the surrounding rock in a withdrawal roadway. However, the fluctuation of mining-induced stress has usually been ignored in previous studies, and it may be the most crucial cause of the significant deformation and instability of the surrounding rock. The thickness of the anchored rock beam is the most vital factor affecting the maximum subsidence and maximum tensile stress of the roof, and increasing the thickness of the anchored rock beam can significantly improve the stability and anti-disturbance performance of the roof. In the proposed double-layer flexible supporting technique, flexible steel strands serve as the carrier, which overcomes the constraint of the roadway height on the length of roof support components. The first layer of flexible support is used to construct a thick fundamental anchorage layer, while the second layer is employed to construct a thicker reinforced anchorage layer, facilitating the effective resistance of the roof against strong mining disturbances. The effectiveness of this technique was further validated through the application of an engineering practice. The research results have reference value for solving the difficult problem of mining roadway support.

Research on the Deformation Law and Stress Distribution Characteristics of Overburden under the Influence of Multiple Faults

Taking the geological conditions of a coal mine in Liupanshui City, Guizhou Province, as the background. The fracture law of the roof in the coal face and the height of the “Two zones” were obtained through theoretical analysis and empirical formula calculation, and the results were verified using similar simulation experiments. Through numerical simulation were researched the stress distribution characteristics of the overlying strata on the working face under the influence of multiple faults. The results showed that in the early stage of mining, F1 undergoes activation under the influence of the main roof weighting, leading to an increase in the degree of plastic damage to the coal seam and strata behind the working face, causing a redistribution of the behind abutment pressure. Finally, under the influence of the F1 barrier effect, the vast majority of the behind abutment pressure is transferred to the lower strata of F1, leading to a continuous increase in stress concentration there. But as the working face continues to advance, the blocking effect of the roof plastic zone above the goaf on the transmission of the behind abutment pressure of the working face begins to appear and continues to increase, leading to a decrease in the peak stress at the lower strata of F1. The stress change process on the F2 side is similar to that on the F1 side, the front abutment pressure is initially distributed in the strata in front of the working face, but as the activation degree of F2 continues to intensify, it gradually shifts to the lower strata of F2, and ultimately the majority of the front abutment pressure is borne by the lower strata of F2.

Study on deflection of principal stress and deformation characteristics of roadway surrounding rock under mining action in close-distance coal seams

The traditional index to determine the width of the stopping coal pillar (WSCP) is generally only the front abutment pressure, but the direction of the principal stress will deflect due to the mining action, which brings significant challenges to controlling the surrounding rock. Based on theoretical calculation and numerical simulation, we studied the deflection law of principal stress under mining action, quantified the bearing capacity of roadway surrounding rock, and determined WSCP. The results show that: (1) With the deflection of the principal stress, the distribution of the surrounding rock plastic zone also rotates; (2) λ and repeated mining dominate and aggravate the deflection angle and range of the principal stress in front of the working face, respectively. When λ is close to 1, the deflection of principal stress should be paid attention to in determining WSCP; (3) The anchorage composite bearing structure (ACBS) provides a scientific basis for quantifying the bearing capacity of surrounding rock. WSCP of the lower coal seam is determined to be 60–75 m, and the corresponding principal stress deflection is about 135°. The control of the surrounding rock in the diagonal direction where the maximum principal stress is concentrated should be paid attention to.

Seismic Reduction Mechanism and Engineering Application of Paste Backfilling Mining in Deep Rock Burst Mines

In the process of coal resources development, a large number of strip coal pillars have been left behind in the coal mines in central–eastern China. With the increase in coal mining depth year by year, the rock burst threat of strip coal pillars is becoming more and more prominent due to the influence of buried depth, geological structure, gob and other factors. Backfilling mining is the main means to recover the residual strip coal pillar. In order to investigate the effect of backfilling mining on the prevention and control of rock burst, taking the paste backfilling workface 1# of Gucheng coal mine as the engineering background, a comprehensive research method of theoretical analysis, numerical simulation and field monitoring was used to study the evolution of stress and of the overburden spatial structure of the backfilling workface under the control of the backfilled ratio. The results showed that the backfilling mining controls the movement and deformation of overburden by reducing the activity range of roof strata. The overburden fracture development height decreases with the increase in backfilled ratio, but there is a boundary effect influenced by the roof deflection before backfilling and the defective distance of roof contact. With the increase in backfilled ratio, the concentration coefficient of front abutment pressure, the vertical displacement of the roof and the development height of the plastic zone of overlying strata decreased obviously, which indicates that filling mining can effectively control the stress of surrounding rock and the movement of overlying strata. The field monitoring data showed that the influence range of the front abutment pressure of the paste backfilling workface was about 90 m and the maximum stress of the surrounding rock of the two entries did not exceed 7 MPa. The average daily frequency of microseism was 1.34, and the average daily total energy of microseism was 1.80 + 103 J, which decreased by 69% and 90%, respectively, compared with the caving method working face with similar geological conditions. The data above showed that the backfilling mining can effectively reduce the working face stress level and dynamic load strength to achieve the effect of prevention and control of rock burst.

Strata pressure behavior and overlying strata movement characteristics of slicing backfill mining in thick coal seams

To improve the coal recovery rate, paste backfill mining was used for 3# coal seam under the industrial square in Panji No.4 Coal Mine. Theoretical analysis, numerical simulation, and similar simulation were applied to study the underground pressure behavior and the overlying strata movement. Results demonstrate the following. (a) Caving mining exerts IV-level mining-induced damages to ground buildings, while slicing backfill mining only causes I-level mining-induced damages to ground buildings. (b) The maximum surface subsidence coefficient of backfill mining is 0.02, and the maximum surface subsidence value is only 4.12% of that under caving mining. (c) The front abutment pressure peak of the working face in caving mining is 28.31 MPa and only 17.82 MPa in backfill mining, which is only 62.95% of caving mining. The distance from the stress peak to the coal wall decreases from 11.14 to 3.04 m. Therefore, the paste backfilling mining of 3# coal seam below the industrial square of Panji No. 4 Coal Mine can effectively control the overlying strata movement and decrease the surface subsidence significantly. Compared with caving mining, backfill mining significantly relieves the strata pressure behaviors and narrows the distance of mining influence.

Research on Roof Load Transfer by Passing Coal Pillar of Working Face in Shallow Buried Closely Multiple-Seam

The dynamic load effect of supports is mainly caused by the movement of the roof structure and the load transfer of overburden. In view of the practice issue that the phenomenon of strong ground pressure is easy to happen, when the working face of the lower coal seam passes the inclined coal pillar in shallow buried closely multiple-seam, it will lead to supprot damaged. This paper takes the mining of over-inclined coal pillars in the 22410 working face of the Bulianta Coal Mine as the background, based on the research method combining the field measurement, physical simulation experiment, and numerical calculation, the evolution law of the front abutment pressure (FAP) and roof weighting in mining under the inclined coal pillar is analyzed, and the mechanism of the stress transfer of the inclined coal pillar and the dynamic load of the support is revealed. The research shows that the concentrated stress of the coal pillar is jointly borne by the front coal wall of the working face and the interburden structure above the support. The vertical stress transmitted from the coal pillar to the floor acts on the key blocks of the interburden of the lower coal seam, which causes strong pressure and dynamic load effect, such as roof structure cut-off. The periodic breaking of the key stratum of the interburden leads to the development height and range of the cracks increasing stepwise. The partition characteristics of the mutual transformation of the interburden stress, the FAP, and the working resistance (WR) by passing the coal pillar stage are revealed, which is divided into three stages and four regions. With the working face passing through the inclined coal pillar, the influence area of the concentrated stress of the coal pillar is reduced, and the peak stress of the coal pillar is gradually transferred to the outside of the coal pillar. When the working face is 5 m away from the coal pillar, the peak of FAP and WR reaches the maximum values, the roof is cutting along the peak stress line, and the working face has a strong weighting phenomenon. The research results are consistent with the field measurement results, providing a reference for the mining of working faces under similar conditions.

Movement Law of Overlying Strata and Abutment Pressure Redistribution Characteristic Based on Rigid Block

Abstract Roof movement induced by coal excavation is the immediate cause of rock pressure redistribution and strata behavior. The rigid block in PFC3D was used to generate a multijointed rock mass, and the PFC3D–FLAC3D coupling model was used to study the movement law of the highly developed structural plane of the overlying strata. Strata movement and abutment pressure redistribution characteristics were obtained. The numerical simulation results showed that the multijointed rock mass model reproduced a rock mass with highly developed structural planes. After coal seam mining, the immediate roof caved and filled the goaf, forming an irregular and regular caved zone. The immediate roof shear slipped along the coal wall. The fracture of the basic roof formed a fractured zone, and the maximum height of the fractured zone first increased and then decreased, exhibiting continuous slow subsidence. The fluctuation of the front abutment pressure was reduced, and the abutment pressure in the goaf jumps was discontinuous. The abutment pressure in the goaf was high in the middle and low on both sides. After the initial fracture of the basic roof, the stress concentration of some rock blocks in the goaf exceeded the in-situ stress, and the average abutment pressure increased with the working face advancing length. With the coal wall of the working face gradually moving away from the goaf, the abutment pressure of the goaf first increased and then remained unchanged; the porosity first decreased sharply and then declined slowly; the coordination number of particles rose sharply and then increased slowly, indicating that the goaf gradually stabilized. Similar simulation results indicated that the variation law of abutment pressure, caving characteristics of the immediate roof, and continuous slow subsidence of the basic roof were the same as those of the numerical simulation.

Study on Evolution of Front Abutment Pressure at Working Face in Repeated Mining of Close-Distance Coal Seams

In order to determine the reasonable width of a stopping coal pillar in close-distance coal seams, the evolution law of front abutment pressure of the working face with repeated mining was studied. Based on the actual engineering project, we conducted field measurement, theoretical analyses, numerical simulations and a physical similarity simulation test to study. The results show that: (1) according to field measurement, the influence range of front abutment pressure increases from 60 m to 75 m with repeated mining; (2) according to theoretical analysis, the arch height and span are negatively and positively correlated with the influence range of front abutment pressure, respectively; (3) with repeated mining, the arch height increased to 165 m, the arch span to 235 m and the influence range to 83.5 m by 14.5 m relative to that before repeated mining; (4) if it is necessary to ensure that the main roadway is less affected by the mining stress, the width of the stopping coal pillar in 2214 working face should be greater than 80 m. The influence range of front abutment pressure increases obviously with repeated mining in close-distance coal seams. The study provides a reference for similar engineering projects.

Comprehensive Monitoring Research on the Effect of Roof Cutting and Pressure Release: A Case Study of the Jiulishan Coal Mine in China

The longwall panel of the Jiulishan coal mine is affected by floor-confined water, while the roof rock weighting makes the production management of some longwall panels difficult. This has caused engineering methods of roof cutting and pressure release to be considered. To explore the dynamic effects of roof cutting and pressure release method, the 14141 longwall panel was taken as the test point, and a variety of monitoring methods were used to comprehensively analyze the characteristics of roof rock weighting, front abutment pressure, and floor mining failure. The main results are as follows: first, the weighting interval, the weighting intensity distribution, and the gateway deformation are asymmetrical. Second, the peak point of the front abutment pressure is approximately 6 m away from the coal wall, the peak stress is 7.02 MPa, the concentration coefficient is 1.54, the range of the pressure increasing area is approximately 16.4 m, and the farthest end of the front abutment pressure effect area is 22.4 m away from the coal wall. Furthermore, the maximum depth of the floor failure is at the No.17 measuring point (A117) position under the condition of one electrode distance. The maximum damage depth is 9.8 m, and the horizontal distance between the location of the maximum failure depth and the working face is 4.1~10.3 m. Finally, compared with the formula calculation value without considering the effect of roof cutting and pressure release or the adjacent 14121 longwall panel, the weighting interval and the weighting intensity, the gateway deformation, the degree of coal wall spalling, the influence range of the front abutment pressure, and the degree of floor mining failure are significantly reduced. The research results provide scientific and technological support for roof cutting and pressure release as a technical method to prevent floor water inrush and reduce the roof rock weighting intensity on the longwall panel.

Mining-Induced Stress and Ground Pressure Behavior Characteristics in Mining a Thick Coal Seam With Hard Roofs

The hard roof of coal mines has the characteristics of high hardness, good integrity, and large layer thickness, which leads to many ground control problems. To reveal the influence of a hard roof structure on the stress performance and coal pillar stability during mining operations, the 8,104 and 8,105 working faces in the Tongxin coal mine were considered as the research object to analyze the stress behavior during the working face advance. Numerical simulation software FLAC3D was used to establish the numerical model of the longwall face under hard roof conditions. The stress distribution laws and coal pillar stability under different roof strengths were analyzed so as to explain the impacts of the hard roof on the stress distribution at the working face. The results show that during the second face proceeding, the influence zone of the front abutment pressure under hard roof conditions is 6 m wider than that under soft roof conditions, and the bearing stress at the working face is 10.4 MPa higher. At the mining position, the plastic zone of the pillar under hard roof conditions is 11 m wider than that under soft roof conditions, and the peak vertical stress is 5.13 MPa higher than that under soft roof conditions. At 25 m ahead of the working face, the plastic zone of the pillar under hard roof conditions is 6 m wider than that under soft roof conditions, and the peak vertical stress is 24.84 MPa higher than that under soft roof conditions. Additional overburden pressure produced by the uncaved hard roof increased pillar stress and plastic zones. Therefore, the hard roof is the main cause of strong ground pressure behavior in the Tongxin coal mine. Aiming at the strong mine pressure behavior, it is suggested to adopt the pre-splitting technology to reduce the influence of the hard roof on mine pressure.

Numerical Simulation of the Stress Field in Repeated Mining of Coal Seams Based on In Situ Stress Measurement

We established an evaluation index of the rock mass stress state for underground coal mines using the strength-stress ratio based on the measured in situ stress and the generalized Hoek–Brown strength criterion. Three in situ stress states, σcm/σ1m < 1.4 (high), 1.4 < σcm/σ1m < 3.6 (medium), and σcm/σ1m > 3.6 (low), were established based on the value of the unconfined compressive strength (σcm) and the maximum principal stress of the rock mass (σ1m). This index classifies the Burtai mine as a medium-high in situ stress field, which is in agreement with the on-site situation, establishing the reliability of the index. The working face was a three-dimensional geological model based on the log sheets. The initial conditions for the model were determined using the combined measurements of the in situ stress regression model. We performed numerical simulations of the roof stress field distribution under repeated mining. Mining the overlying coal seam leads to significant variation in the value and direction of the main roof, σ1, within the range of the front abutment pressure under the pillar and gob. Along the main roof strike direction, σ1 under the pillar is 1.5 times that under the gob, and the σ1 direction under the pillar is deflected by 5°, which is 30° smaller than that under the gob. This provides a reference for optimized underground coal mining.

Stress and deformation analysis of gob-side pre-backfill driving procedure of longwall mining: a case study

At present, non-pillar entry protection in longwall mining is mainly achieved through either the gob-side entry retaining (GER) procedure or the gob-side entry driving (GED) procedure. The GER procedure leads to difficulties in maintaining the roadway in mining both the previous and current panels. A narrow coal pillar about 5–7 m must be left in the GED procedure; therefore, it causes permanent loss of some coal. The gob-side pre-backfill driving (GPD) procedure effectively removes the wasting of coal resources that exists in the GED procedure and finds an alternative way to handle the roadway maintenance problem that exists in the GER procedure. The FLAC3D software was used to numerically investigate the stress and deformation distributions and failure of the rock mass surrounding the previous and current panel roadways during each stage of the GPD procedure which requires "twice excavation and mining". The results show that the stress distribution is slightly asymmetric around the previous panel roadway after the “primary excavation”. The stronger and stiffer backfill compared to the coal turned out to be the main bearing body of the previous panel roadway during the "primary mining". The highest vertical stresses of 32.6 and 23.1 MPa, compared to the in-situ stress of 10.5 MPa, appeared in the backfill wall and coal seam, respectively. After the "primary mining", the peak vertical stress under the coal seam at the floor level was slightly higher (18.1 MPa) than that under the backfill (17.8 MPa). After the "secondary excavation", the peak vertical stress under the coal seam at the floor level was slightly lower (18.7 MPa) than that under the backfill (19.8 MPa); the maximum floor heave and maximum roof sag of the current panel roadway were 252.9 and 322.1 mm, respectively. During the "secondary mining", the stress distribution in the rock mass surrounding the current panel roadway was mainly affected by the superposition of the front abutment pressure from the current panel and the side abutment pressure from the previous panel. The floor heave of the current panel roadway reached a maximum of 321.8 mm at 5 m ahead of the working face; the roof sag increased to 828.4 mm at the working face. The peak abutment pressure appeared alternately in the backfill and the coal seam during the whole procedure of "twice excavation and mining" of the GPD procedure. The backfill provided strong bearing capacity during all stages of the GPD procedure and exhibited reliable support for the roadway. The results provide scientific insight for engineering practice of the GPD procedure.

Analysis on Rock Burst Risks and Prevention of a 54 m-Wide Coal Pillar for Roadway Protection in a Fully Mechanized Top-Coal Caving Face

A case study based on the 401103 fully mechanized caving face in the Hujiahe Coal Mine was carried out in this research to analyze the rock burst risks in a 54 m-wide coal pillar for roadway protection. Influencing factors of rock burst risks on the working face were analyzed. Stress distribution characteristics on the working face of the wide coal pillar for roadway protection were discussed using FLAC3D numerical simulation software. Spatial distribution characteristics of historical impact events on the working face were also investigated using the microseismic monitoring method. Results show that mining depth, geological structure, outburst proneness of coal strata, roof strata structure, adjacent mining area, and mining influence of the current working face are the main influencing factors of rock burst on the working face. Owing to the collaborative effects of front abutment pressure of the working face and lateral abutment pressure in the goaf, the coal pillar is in the ultimate equilibrium state and microseismic events mainly concentrate in places surrounding the coal pillars. Hence, wide coal pillars become the regions with rock burst risks on the working face. The working face adopts some local prevention technologies, such as pressure relief through presplitting blasting in roof, pressure relief through large-diameter pores in coal seam, coal seam water injection, pressure relief through large-diameter pores at bottom corners, and pressure relief through blasting at bottom corners. Moreover, some regional prevention technologies were proposed for narrow coal pillar for roadway protection, including gob-side entry, layer mining, and fully mechanized top-coal caving face with premining top layer.

The fracturing models of hard roofs and spatiotemporal law of mining-induced stress in a top coal caving face with an extra-thick coal seam

Hard roof fracturing characteristics and mining-induced stress behaviors can be studied to ensure production safety in top coal caving face. Field tests were conducted at top coal caving face No. 8309 in the Tongxin mine of the Datong Coal Mining Group in China. The monitoring results indicated two modes of periodic weighting: short weighting steps with low intensity and long weighting steps with high intensity. According to three periodic weighting phenomena, three models of hard roof fracturing were proposed based on the cantilever-voussoir beam compound structure. And the equations of the working resistance for long/high periodic weighting were derived. Furthermore, a difference in the pressure changes among the front leg, back leg and front beam of the supports was observed, and the back leg pressure significantly increased during periodic weighting. The change in the single prop pressure was divided into a slow growth stage and a rapid growth stage. And the front abutment pressure fluctuated with the periodic instability of the roof compound structure. Moreover, the fracturing of the roof above the stope resulted in roadway deformation. The instability of the cantilever-voussoir beam structure also intensified the roadway deformation. Finally, a support-surrounding rock mechanical model was established considering the behavior of an elastic cantilever beam. The required work resistance considering the bearing capacity of top coal were discussed, providing a basis for reasonable support selection. These research results can provide references for mining pressure and stratum movement control.

Study on Floor Mechanical Failure Characteristics and Stress Evolution in Double Predriven Recovery Rooms

This paper takes the double predriven recovery rooms (DPRR) of 31109 panel of a coal mine in Inner Mongolia as a case study. DPRRs are used to withdraw mining equipment, which play a significant role in safe and efficient production in the final longwall mining stage. Theoretical analysis and numerical simulation were carried out to study the reasonable size of the front abutment pillar between DPRR (inter-DPRR pillar) and the damage depth of the DPRR floor. The results show that (1) the stress distribution of the fender (the remnant longwall panel) can be approximately divided into three stages with the advance of the working face: stress redistribution (the first) stage, stress superimposed growth (the second) stage, and stress transfer (the third) stage. (2) According to stress distribution and the corresponding failure mode of the fender, the calculation model of the slippage damage of the DPRR floor is rectified, and the damage range of the floor is rezoned to make it more suitable for the damage depth of the room. (3) The zone of influence of the front abutment pressure is 40–50 m, and the stress around the DPRR increases significantly in the final mining stage. When the size of the inter-DPRR pillar is greater than 15 m, the effect of increasing the coal pillar size on lowering the peak stress of the main predriven recovery room is limited. (4) Floor heave tends to increase at first and then decrease with depth and reaches the maximum in the depth of 5 m in the final mining stage, indicating that 5 m is the starting point for the initial depth of the floor heave. (5) The theoretical calculation shows that the reasonable size of the inter-DPRR pillar is 20 m, and the critical width of the fender is 18.48 m, which can guide the secondary support to prevent dynamic disasters. Floor grouting and constructing concrete floor are effective and economic ways to control the floor heave.