• Geomechanics and Engineering
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• v.13 no.6
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• pp.997-1010
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• 2017

The black box model is a relatively new option for nonlinear dynamic system identification. It can be used for prediction problems just based on analyzing the input and output data without considering the changes of the internal structure. In this paper, a black box model was presented to solve unconstrained overlying strata movement problems in coal mine production. Based on the black box theory, the overlying strata regional system was viewed as a "black box", and the black box model on overburden strata movement was established. Then, the rock mechanical properties and the mining thickness and mined-out section area were selected as the subject and object respectively, and the influences of coal mining on the overburden regional system were discussed. Finally, a corrected method for height prediction of the fractured zone was obtained. According to actual mine geological conditions, the measured geological data were introduced into the black box model of overlying strata movement for height calculation, and the fractured zone height was determined as 40.36 m, which was comparable to the actual height value (43.91 m) of the fractured zone detected by Double-block Leak Hunting in Drill. By comparing the calculation result and actual surface subsidence value, it can be concluded that the proposed model is adaptable for height prediction of the fractured zone.

• Geomechanics and Engineering
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• v.21 no.3
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• pp.301-313
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• 2020

In the process of mining closely spaced coal seams, the problem of roadway arrangement in lower coal seams has long been a concern. By means of mechanical model calculation and numerical simulation postprocessing, the distribution of the stress deviator below the floor of a goaf and the evolution of the stress deviator in the vertical and horizontal directions are studied under the influence of horizontal stress. The results of this theoretical study and numerical simulation show that the stress deviator decreases exponentially with increasing depth from the floor below the coal side. With the increase in the horizontal stress coefficient λ, the stress deviator concentration area shifts. The stress deviator is concentrated within 10 m below the goaf and 15 m laterally from the coal side; thus, the magnitude of the surrounding rock stress deviator should be considered when planning the construction of a roadway in this area.

• Geomechanics and Engineering
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• v.17 no.3
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• pp.295-305
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• 2019

The bedding plane has a significant influence on the effect of blasting fragmentation and the overall performance of underground mining. This paper explores the effects of fragmentation of the bedding plane and different angles by using the numerical analysis. ANSYS/LS-DYNA code was used for the implementation of the models. The models include a dynamic compressive and tensile failure which is applied to simulate the fractures generated by the explosion. Firstly, the cracks propagation with the non-bedding plane in the coal with two boreholes detonated simultaneously is calculated and the particle velocity and maximum principal stress at different points from the borehole are also discussed. Secondly, different delay times between the two boreholes are calculated to explore its effects on the propagation of the fractures. The results indicate that the coal around the right borehole is broken more fully and the range of the cracks propagation expanded with the delay time increases. The peak particle velocity decreases first and then increases with the distance from the right borehole increasing. Thirdly, different angles between the bedding plane and the centerline of the two boreholes and the transmission coefficient of stress wave at a bedding plane are considered. The results indicated that with the angles increase, the number of the fractures decreases while the transmission coefficient increases.

• Geomechanics and Engineering
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• v.29 no.4
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• pp.451-461
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• 2022

The roadways surrounded by rock and coal will lose their stability or even collapse under rock burst. Rock burst mainly involves an evolution of dynamic loading which behaves quite differently from static or quasi-static loading. To compare the dynamic response of coal and rocks with different static strengths, three different rocks and bituminous coal were selected for testing at three different dynamic loadings. It's found that the dynamic compression strength of rocks and bituminous coal is much greater than the static compression strength. The dynamic compression strength and dynamic increase factor of the rocks both increase linearly with the increase of the strain rate, while those of the bituminous coal are irregular due to the characteristics of multi-fracture and heterogeneity. Moreover, the absorbed energy of the rocks and bituminous coal both increase linearly with an increase in the strain rate. And the ratio of absorbed energy to the total energy of bituminous coal is greater than that of rocks. With the increase of dynamic loading, the failure degree of the sample increases, with the increase of the static compressive strength, the damage degree also increases. The static compassion strength of the bituminous coal is lower than that of rocks, so the number of small-scale fragments was the largest after bituminous coal rupture.

• International journal of steel structures
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• v.18 no.5
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• pp.1801-1817
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• 2018

This paper presents the numerical parametric analysis on the loading capacity of Channel purlins with screw-fastened sheeting, in which the effects of anti-sag bar and corrugated steel sheet on the ultimate capacity are studied. Results show that the setup of anti-sag bars can reduce the deformations and improve the ultimate capacity of C purlins. The traditional method of setting the anti-sag bars in the middle of the web is favorable. The changing of sheeting type, sheeting thickness and rib spacing has significant effects on the ultimate capacity of C purlins without anti-sag bars, compared with those with anti-sag bars. The proposed design formulas are relatively consistent with the calculations of EN 1993-1-3:2006, which is different from those of GB 50018-2002.

• Geomechanics and Engineering
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• v.36 no.5
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• pp.427-440
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• 2024

This study concentrates on the 301 comprehensive caving working face, notable for its considerable mining height. The roof model is established by integrating prior geological data and the latest borehole rock stratum's physical and mechanical parameters. This comprehensive approach enables the determination of lithology, thickness, and mechanical properties of the roof within 50 m of the primary mining coal seam. Utilizing the transfer rock beam theory and incorporating mining pressure monitoring data, the study delves into the geometric parameters of the direct roof, basic roof movement, and roof pressure during the initial mining process of the 301 comprehensive caving working face. The direct roof of the mining working face is stratified into upper and lower sections. The lower direct roof consists of 6.0 m thick coarse sandstone, while the upper direct roof comprises 9.2 m coarse sandstone, 2.6 m sandy mudstone, and 2.8 m medium sandstone. The basic roof stratum, totaling 22.1 m in thickness, includes layers such as silty sand, medium sandstone, sandy mudstone, and coal. The first pressure step of the basic roof is 61.6 m, with theoretical research indicating a maximum roof pressure of 1.62 MPa during periodic pressure. Extensive simulations and analyses of roof subsidence and advanced abutment pressure under varying working face lengths. Optimal roof control effect is observed when the mining face length falls within the range of 140 m-155 m. This study holds significance as it optimizes the working face length in thick coal seams, enhancing safety and efficiency in coal mining operations.

• Geomechanics and Engineering
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• v.22 no.3
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• pp.255-264
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• 2020

With the ongoing development of deep mining of coal resources, some coal mine dynamic disasters have exhibited characteristics of both coal-gas outbursts and rockbursts. Therefore, research is required on the mechanism of rockburst-outburst coupling disaster. In this study, the failure characteristics of coal-rock combination structures were investigated using lab-scale physical simulation experiments. The energy criterion of the rockburst-outburst coupling disaster was obtained, and the mechanism of the disaster induced by the gas-solid coupling instability of the coal-rock combination structure was determined. The experimental results indicate that the damage of the coal-rock structure is significantly different from that of a coal body. The influence of the coal-rock structure should be considered in the study of rockburst-outburst coupling disaster. The deformation degree of the roof is controlled by the more significant main role of the gas pressure and the difference in the strength between the rock body and the coal body. The outburst holes and spall characteristics of the coal body after the failure of the coal-rock structure are strongly affected by the difference in strength between the roof and the coal body. The research results provide an in-depth understanding of the mechanism of rockburst-outburst coupling disasters in deep mining.

• Geomechanics and Engineering
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• v.14 no.5
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• pp.421-428
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• 2018

In deep mining, the lateral deformation of strip coal pillar appears to be a new characteristic. In order to study the lateral deformation of coal-mass, a monitoring method and monitoring instrument were designed to investigate the lateral deformation of strip coal pillar in Tangkou Coalmine with the mining depth of over 1000 m. Because of without influence of repeated mining, the bedding sandstone roof is easy to break and the angle between maximum horizontal stress and the roadway is small, the maximum lateral deformation is only about 287 mm lower than the other pillars in the same coalmine. In deep mining, the energy accumulation and release cause a discontinuous damage in the heterogeneous coal-mass, and the lateral deformation of coal pillar shows discontinuity, step and mutation characters. These coal-masses not only show a higher plasticity but also the high brittleness at the same time, and its burst tendency is more obvious. According to the monitoring results and theoretical calculations, the yield zone of the coal pillar width is determined as 15.6 m. The monitoring results presented through this study are of great significance to the stability analysis and design of coal pillar.

• Geomechanics and Engineering
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• v.19 no.2
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• pp.167-178
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• 2019

Open-pit (OP) and underground (UG) mining are usually used to exploit shallow and deep ore deposits, respectively. When mine deposit starts from shallow subsurface and extends to a great depth, sequential use of OP and UG mining is an efficient and economical way to maintain mining productivity. However, a transition from OP to UG mining could induce significant rock movements that cause the slope instability of the open pit. Based on Yanqianshan Iron Mine, which was in the transition from OP to UG mining, a large-scale two-dimensional (2D) model test was built according to the similar theory. Thereafter, the UG mining was carried out to mimic the process of transition from OP to UG mining to disclose the triggered rock movement as well as to assess the associated slope instability. By jointly using three-dimensional (3D) laser scanning, distributed fiber optics, and digital photogrammetry measurement, the deformations, movements and strains of the rock slope during mining were monitored. The obtained data showed that the transition from OP to UG mining led to significant slope movements and deformations that can trigger catastrophic slope failure. The progressive movement of the slope could be divided into three stages: onset of micro-fracture, propagation of tensile cracks, and the overturning and/or sliding of slopes. The failure mode depended on the orientation of structural joints of the rock mass as well as the formation of tension cracks. This study also proved that these non-contact monitoring technologies were valid methods to acquire the interior strain and external deformation with high precision.

• Geomechanics and Engineering
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• v.12 no.3
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• pp.541-560
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• 2017

In this research, experimental and numerical simulations were adopted to investigate the effects of ligament angle on compressive strength and failure mode of rock-like material specimens containing two non-coplanar filled fissures under uniaxial compression. The experimental results show that with the increase of ligament angle, the compressive strength decreases to a nadir at the ligament angle of $60^{\circ}$, before increasing to the maximum at the ligament angle of $120^{\circ}$, while the elastic modulus is not obviously related to the ligament angle. The shear coalescence type easily occurred when ${\alpha}$ < ${\beta}$, although having the same degree difference between the angle of ligament and fissure. Numerical simulations using $PFC^{2D}$ were performed for flawed specimens under uniaxial compression, and the results are in good consistency with the experimental results. By analyzing the crack evolution process and parallel bond force field of rock-like material specimen containing two non-coplanar filled fissures, we can conclude that the coalescence and propagation of crack are mainly derived from parallel bond force, and the crack initiation and propagation also affect the distribution of parallel bond force. Finally, the displacement vectors in ligament region were used to identify the type of coalescence, and the results coincided with that obtained by analyzing parallel bond force field. These experimental and numerical results are expected to improve the understanding of the mechanism of flawed rock engineering structures.