Search

A considerable number of coal mines employed room and pillar mining in the last century in northern China, where the goaf remained stable for a period of time; however, with the increased exposure of coal pillars, their collapse may gradually increase. The stability assessment of these old rooms and pillar goafs is challenging due to their concealment, irregular mining patterns, and the long passage of time. The methodology developed in this study, based on "space-sky-earth" remote sensing such as InSAR to trace historical deformation, the UAV observation of current surface damage, and comparison of mining spaces, can rapidly detect on a large scale the collapse of old goafs and the trend of damage. This study is conducted with an example of a coal mine in Yulin, Northern China, where obtained quantitative surface deformation values were integrated with qualitative surface damage interpretation results, followed by a yearly analysis of the overlying rock movement in accordance with the underground coal mining process. The results show that from 2007 to 2021, corresponding surface deformation and damage occurred following mining progress. However, the room and pillar goaf areas had not undergone any surface deformation, nor had there been incidents of landslides or ground fissures; therefore, it was speculated that no roof collapse had occurred in this region. The surface deformation and damage associated with underground coal mining are complex and influenced by the coal seam occurrence, mining methods, strata lithology, terrain slope, temporal evolution, and anthropogenic modifications. These phenomena are representative of the coal mining area, and this methodology can provide a reference for similar endeavors. [ABSTRACT FROM AUTHOR]

Traditional coal mining processes have increasingly raised concerns about roadways and surface damage. An innovative mining method, based on DSRC technology, has been proposed to address these issues. The DSRC technology involves using bilateral energy gathering devices to achieve roof cutting on both sides of the working face. The N00 method builds upon DSRC technology by incorporating additional auxiliary techniques to achieve roadways and surface protection during mining. With this method, mining N working faces can be accomplished with the advancement of 0 roadway and the retention of 0 coal pillar. The protection mechanism of the DSRC method on roadways and surface is analyzed through theoretical calculation, numerical simulation, and field engineering analysis. The N00 method with DSRC modifies the structure of the roadway roof by changing the long cantilever beam to a short cantilever beam. This modification reduces the stress on the roof and improves the stability of the roadway roof. The DSRC cuts off the stress transmission in a part of the roadway roof, reducing stress concentration on the coal side and enhancing its stability. The N00 method with DSRC increases the amount of strata bulking, utilizing bulking gangue to support the overlying strata, reducing the development height of the plastic zone in the gob, protecting the overlying aquifers, and mitigating damage to the surface ecological environment. Field monitoring results show that the N00 method with DSRC reduces the door-like support stress, NPR anchor cable stress, and deformation of the roadway, decreases stress concentration on the coal side, minimizes surface deformation, and reduces water inflow into the working face. Therefore, the purpose of protecting roadways and the surface is achieved. Highlights: Double-sided roof cutting technique has been proposed to protect roadways and the surface strata. Roadways are protected by altering the structure of the roof and cutting off part of the stress transmission. Surface strata are protected by increasing the amount of strata bulking. [ABSTRACT FROM AUTHOR]

At present, when addressing the problem of fault stability in mines, most attention is paid to the study of the impact of a single mining method on faults. In order to study fault stability in mines more comprehensively, this paper researches the effect of coordinated mining using multiple mining methods. For example, the sublevel caving method without the sill pillar and the lower-layer filling method of coordinated mining can be used to analyze the dynamic response law of the fault, as well as the stability of the fault in different mining conditions. In this paper, the stress field, displacement field and sliding trend index of the fault plane and orebody near the fault are obtained via numerical simulation and theoretical analysis methods, and the main factors affecting fault stability under different mining conditions are analyzed. The results show that under the influence of cooperative mining, the shear stress fluctuation of the fault surface and the ore body near the fault increase gradually with the sublevel height of the sublevel caving method without the sill pillar and the lower-layer filling method, and the indexes of slip tendency become larger, which may be a precursor of fault activation. In particular, the fault surface of the sublevel caving method without the sill pillar reflects the drastic change in the shear stress with the larger displacement, and the instability brought to the fault by the sublevel caving method without the sill pillar is greater than that brought to the fault by the lower layer filling method under the premise of only changing the height of the sublevel. [ABSTRACT FROM AUTHOR]

Sustainable development in coal mining requires a continuous and efficient method of coal extraction. Research shows that gob-side entries retained through roof cutting retained gob-side (RCGE) are vital for improving mining efficiency, enhancing coal recovery rates, and enabling continuous production. However, the mechanism of surrounding rock deformation during close-distance co-mining of coal seams with this technique is not yet clear. For the Jiaokou coal mine in China, due to an unreasonable stagger distance between upper and lower working faces, the gob-side entries retained at the 9102 tailgate and 10102 headgate experience severe rock pressure, leading to significant prop damage and a sharp reduction in the cross-section of the entry. This greatly hampers the reuse of these entries. To investigate this issue, we established a model to study the stress distribution of surrounding rocks at different stagger distances (20 m, 40 m, 60 m, 80 m, and 120 m) through numerical simulation and optimized the support parameters for the retained entries. Our research found that when the subsidence of the roof in the upper coal seam exceeds 0.74 m but is less than 1.33 m, there is sliding instability in the mining body. When the subsidence exceeds 1.33 m, the mining body will rotate and deform, causing significant mining pressure within the retained entry. A stagger distance of 40 m between the upper and lower working faces can reduce pressure on the face during the mining of the lower coal seam. Extensive field measurements of rock pressure revealed that the damage rate of the single column in the gob-side entries of the upper and lower coal seams does not exceed 5% and 1%, respectively. In summary, this study provides a practical method to reduce damage to entries during the mining process, thereby increasing the continuous production capability of the coal mine. This is critical for the sustainable development of coal mining. [ABSTRACT FROM AUTHOR]

In longwall top coal caving (LTCC) mining methods, problems such as large deformations, loss of coal resources, and excavation relay out of balance are often encountered with the increased production in single mining panels. To address these problems, a novel non‐pillar coal mining technology in LTCC was used based on the equilibrium mining theory. This approach could cut off the stress transmission, eliminate the coal pillar, and retain the gob‐side entry. In this technology, the entry is first excavated along the roof of the coal, and then four techniques are applied to control the surrounding rock of the retained entry. The constant resistance and large deformation anchor cable are used to maintain the roof stability. The directional roof split blasting technique is adopted to separate the roof between the entry and gob. The blocking‐gangue support system (BGSS) is used to form the gangue rib lagging the mining panel. The temporary support system is applied to prevent the roadway from being affected by the dynamic pressure and prevented the BGSS from falling. Finally, the field monitoring results indicate that the surrounding rock of retained entry is stabilized at 200 m behind the working face. The deformation of retained entry is within a reasonable range at 180–190 m behind the mining panel. The results provide an important reference for the thick coal seam non‐pillar coal mining technology, specifically using this technology in LTCC panels. [ABSTRACT FROM AUTHOR]

Aiming at the problem of difficult mining of coal resources under buildings, railroads and water bodies in northern Shaanxi, China, a narrow strip filling mining (NSFM) method is proposed. The four-wheel cyclic mining and filling technology is systematically introduced, the method of theoretical analysis, numerical simulation and physical similarity simulation experiment is adopted to carry out the research, and the pressure rule of filling body—coal pillar and the control effect of roof in the mining and filling process are clarified. The finite beam model of elastic foundation is established, the discriminative basis of roof instability is put forward, and the key parameters of filling mining are deduced. The numerical simulation results show that in the process of four-wheel cyclic mining and filling, the backfill filled in this stage starts to participate in the load bearing after the next stage of coal pillar is mined, and the backfill regularly shares the overburden load in stages. The overall subsidence pattern of the immediate roof is similar to that of the "square basin", and the final subsidence of the immediate roof is 0.08 m, which is at a relatively low level. The physical similarity simulation experiment verifies the above conclusion, no separation layer and cracks appear in the roof of coal seam, and the filling body and coal pillar maintain stable safe bearing state. This method can improve the recovery rate of coal resources, control the roof subsidence in a small range, and provide a feasible solution for sustainable mining of coal seam under buildings, railroads and water bodies. [ABSTRACT FROM AUTHOR]

Copyright of Nonferrous Metals (Mining Section) is the property of Beijing Research Institute of Mining & Metallurgy Technology Group and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

With the increasing proportion of close-distance coal seam mining in China, the problem of strong mining pressure during the mining of close-distance coal seams is becoming more and more severe. This article focuses on the complex stress environment and severe mining pressure encountered in the mining of thick coal seams under the multi-coal-seam goaf of Zhunnan Coal Mine. By using research methods, such as similar material simulation, theoretical analysis, and numerical simulation, it studies in depth the instability characteristics of the overlying rock structure of the W1701 working face, the inducing factors and mechanisms of strong mining pressure during the mining process, and control measures. The results show that the roof structure of the W1701 working face can be divided into "high-level key layer (hard rock)–giant thick soft and weak rock group–low-level key layer (hard rock)", and the law of mining pressure manifestation presents a small cycle formed by the instability of "masonry beam" structure and a main large cycle formed by the periodic penetration and step-down of the giant thick soft and weak rock group, with the load on the support during the large cycle up to 5.4 times the rated working resistance. In addition, this article proposes the strategy of using layered mining to control the manifestation of strong mining pressure under the "hard sandwiched soft" overlying rock condition of the Zhunnan Coal Mine, optimizes the thickness of the layered mining of the thick coal seam, and finally, determines the upper layer thickness of 2.8 m and the lower layer thickness of 4 m, inducing the giant thick soft and weak rock formation to undergo incremental damage and releasing the fracture energy incrementally, effectively controlling the manifestation threat of strong mining pressure in the mining of thick coal seams under the close-distance coal seam goaf. As the proportion of close-range coal seam mining increases in China, the problem of strong mining pressure during the mining of close-range coal seams becomes more severe. This article focuses on the complex stress environment and severe mining pressure in the mining of thick coal seams under multiple mined-out areas in the Zhunnan coal mine. Similar material simulation, theoretical analysis, and numerical simulation methods were used to conduct in-depth research on the unstable characteristics of the overlying rock structure of the W1701 working face, the causes and mechanisms of strong mining pressure during the mining process, and control measures. The results show that the roof structure of the W1701 working face can be divided into "high-level key layer (hard rock)–thick soft weak rock group–low-level key layer (hard rock)." The law of mining pressure manifestation presents small cycles of instability formed by "block beams" and main cycles of pressure formed by vertically cracked periodic penetration and step sinking of the thick soft weak rock group. Moreover, during the main cycle of pressure, the load-bearing capacity of the support is up to 5.4 times the rated working resistance. Furthermore, it is proposed to use hierarchical mining to control the manifestation of strong mining pressure in the "hard-inlaid soft" overlying rock condition of the Zhunnan coal mine and optimize the thickness of layered mining of thick coal seams. Ultimately, the upper layer thickness was determined as 2.8 m; the lower layer thickness was determined as 4 m, and the layered mining induced the thick soft weak rock group to undergo gradual damage and energy release, effectively controlling the threat of severe mining pressure during the mining of thick coal seams under the close-range coal seam mining. [ABSTRACT FROM AUTHOR]

Copyright of Industrial Minerals & Processing / Huagong Kuangwu yu Jiagong is the property of Industrial Minerals & Processing Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

The re-mined face using top coal caving system is the most applicable method for recovering the remaining reserves of a previous partially-mined thick coal seam. However, this mining method may encounter the problems of low recovery and unpredicted geological conditions. A numerical model is developed using PFC2D for studying the movement of top coal mass and development of coal-rock mass interface at a longwall top coal caving re-mined face. The re-mined face advances in the lower seam below the upper solid coal pillar, previous entries and gob. A theoretical analysis according to the unsteady flow model is developed to calculate the proper time duration of caving operation. The results showed that the top coal to be recovered through the caving window before the initiation of caving operation is a partial spheroid-shaped geometry. The boundary between the coal and rock mass then develops to a funnel-shaped coal-roof interface as the caving operation continues. The top coal recovery is 98.1%, 77.1% and 70.5% for caving operations below the solid coal, entries and gob area in the upper seam, respectively. The proper timing of caving and interval of caving operation is important for obtaining a high coal recovery. Good agreement is achieved between the proposed model and the improved Boundary-Release model for short of B-R model. The study in this work may provide reference for the safety and efficiency of the extraction of the longwall top coal caving re-mined face. [ABSTRACT FROM AUTHOR]

The prediction and control of roof water disaster has been one of the key problems in the mining process of coal resources. With the increase of mining depth in the western mining area, the roof caving and communicating to the overlying water-bearing strata have led to an increasing number of roof water inrush accidents and deterioration of production environment of the working face. Aiming at the problem of roof water disaster prediction of sandy-argillaceous structural strata in Shanghaimiao mining area in China, firstly, the mechanical structure model of roof water inrush was built according to the parameters in the advance of the working face, and the thickness of the sandstone was used as the main controlling geological factor; the formula was derived for calculating the water-rich intensity F zh of the overlying strata in the mining area. Secondly, starting from the height of the "Breakage-arch" development disturbance by combining the mechanical structure model of roof water inrush, the relative positional relation of the "Breakage-arch" and the water-bearing strata was analyzed, and a new method for judging the risk coefficient T W of roof water inrush was proposed. Finally, according to the geological drilling histogram and the field conditions of #111084 working face of the no. 1 mine in Shanghaimiao mine area, the water-rich index and water inrush risk coefficient of overlying water-bearing strata were obtained by means of discriminant analysis, and the water inrush risk of #111084 working face was evaluated and predicted quantitatively. The water-rich property evaluation of water-bearing strata under the condition of low degree of hydrogeological exploration was evaluated accurately and reasonably, and a new evaluation method of the water disaster during mining was proposed. [ABSTRACT FROM AUTHOR]

The longwall mining method with gob-side entry retaining via roof cutting is a new underground coal mining method which has the characteristics of a high resource recovery ratio and environmental friendliness. Due to the complexity of this method, the research method of case-based dynamic on-site monitoring, analysis, adjustment, and optimization is usually adopted. Based on a roadway retaining via roof cutting project, in addition to the traditional indirect monitoring method of hydraulic support pressure, this study innovatively establishes a direct monitoring method for roof caving by monitoring the gangue pressure in the goaf, which provides data for the roof cutting effect and offers a new method for studying the overlying strata movement. In the project, a comprehensive monitoring and analysis system was established, including gangue pressure, cable bolt stress, bracket pressure, roadway deformation, and roof separation, which was used to dynamically analyze the effect of roof cutting and optimize the support design. The results show that the pressure of the hydraulic support close to the roof cutting is low, indicating that roof cutting is favorable in the roadway retaining mining method. The roadway deformation in the advanced abutment pressure area of the working face is small. The mining-induced stress caused by the collapse and compaction of the overlying strata in the goaf is the dominant factor affecting the effect of roadway retaining, especially in the 50–100 m range behind the working face, where the dynamic load causes high bearing capacity of the support elements, large roadway convergence, and roof separation. Temporary support and supplementary reinforcement should be added when necessary. The monitoring system presented in this study is highly comprehensive, simple, reliable, and low in cost, providing a reference for roof cutting roadway retaining projects and roof caving-related studies. [ABSTRACT FROM AUTHOR]

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Underground longwall mining can seriously disturb the surrounding rock masses above the panel. A surface cracking zone, continuous deformation zone, fractured zone, and caved zone can be formed in the overlying strata (termed the "four zones"), which may further result in the spontaneous combustion of coal seams and water inrush. It is essential to study and predict the development characteristics of the four zones induced by longwall mining to guarantee mining safety. These four zones are developed during the mining process, and the mechanical properties in different regions correspondingly differ. Thus, the dynamic zoning characteristics of the disturbed rock masses should be considered in any simulations. In this paper, an FDM-based dynamic zoning method for disturbed rock masses above a longwall mining panel is proposed. This method is mainly composed of four stages: (1) establishing a simplified complete stress-strain curve; (2) determining the zoning criteria; (3) adaptively adjusting the mechanical parameters of the disturbed rock mass; and (4) numerically modeling the longwall mining based on the FDM. The proposed method was applied to a study site in the Taixi coal mine. The dynamic development process of the four zones induced by longwall mining was clearly observed in the modeling procedure. The numerical modeling results achieved in this work, including the periodical coal-seam roof caving and the dynamic development characteristics of the four zones, are consistent with the observed distribution and other studies. The heights of the caved and fractured zones are basically consistent with the empirical formula. Thus, the dynamic zoning method can analyze and predict the dynamic development characteristics of four regions. [ABSTRACT FROM AUTHOR]

Coal mining inevitably results in the movement of overlying strata, with the upward formation of the strata leading to surface subsidence, causing irreversible impact on the buildings, land, and ecological environment. The movement and deformation of the strata are controlled by the bearing structure in the overlying strata, whose failure results in the deformation and breakage of the overlying strata simultaneously. While studies have been conducted on the arch structure in unconsolidated layers (ASUL), its bearing performance has not been addressed. Therefore, this study develops a bearing mechanics model based on the morphological characteristics of the ASUL. The analytical expressions of the axial force, bending moment, and shear force of the cross-sectional area were determined using theoretical derivations. The model analysed the internal forces and showed the influence laws of the overlying load, horizontal pressure coefficient, and rise-to-span ratio of the ASUL. The failure criterion of the bearing was also further determined. The results indicated that with overlying and horizontal loads, the axial force and bending moment are symmetrically distributed, whereas the shear force is asymmetrically distributed. In addition, the axial force gradually increases from the dome to the base of the ASUL. Compared to the axial force and bending moment, the shear force has a lower impact on the stability of the ASUL. Most of the axial force and overlying load is received through the axial compression of the cross-section to maintain stability and play a bearing role on the overlying unconsolidated layers. As the overlying load, horizontal pressure coefficient, and rise-to-span ratio increase, the axial force, bending moment, and shearing force also increase gradually. This effect is more apparent at the dome, spandrel, and base of the ASUL. The stability of the dome and spandrel is key to the overall structural stability. Therefore, the failure criterion for the ASUL was determined based on the compression failure at the dome and spandrel. During the mining process of the working face, the ASUL served as load-bearing control for the overlying unconsolidated layers. Further, increasing width of the working face damages and shifts the base of the ASUL, resulting in compression failure at the dome and spandrel, further inducing dome lift and causing overall failure of the ASUL. Considering the aforementioned factors, a control method that reinforces the surface subsidence of the ASUL by 'one-time, upward, staged, and multiple-ground-drilling' compaction grouting has been proposed. During the mining process of the working face, the arch bead-like structure, combined with the ASUL, serves as the load-bearing control on the overlying strata and ground surface, reducing ASUL deformation in the unconsolidated layers, overlying strata, and ground surface. This process enables the controlling of ground subsidence of coal mining in thick unconsolidated layers. [ABSTRACT FROM AUTHOR]

The overlying weakly cemented, and poorly performing strata in Jurassic mines of western China have mechanical properties that are generally lower than those in the Carboniferous-Permian coal mines of central and east China. During coal mining, the overlying strata easily deform and fracture. These then formed water-conducting channels, triggering a series of eco-environmental issues, including ground fracturing, collapsed surfaces, declined underground water levels, deserted lands, and even severe water/sand burst accidents. To study the fracture characteristics of weakly cemented overlying strata and the evolution law of water-conducting fractures in Jurassic coal mines in western China, this study selected Tashidian Erjingtian Mine in Korla, Xinjiang, as the research object. Based on the simulation data obtained with physical analog model testing and field monitoring results, the authors investigated the development of water-conducting fractures in the weakly cemented overlying strata during the coal seam mining process. We simultaneously determined the location of key strata in the working face based on key stratum theory. According to the present research results, key strata controlled the development height of water-conducting fractures. When the primary key stratum or sub-key stratum was not fractured, the development of water-conducting fractures was stagnant; water-conducting fractures developed abruptly when the primary key stratum or sub-key stratum was cracked. The heights of water-conducting fractures in the weakly cemented overlying strata of western China exceeded that of similar stopes of central and east China. These research results provided theoretical and technical support for safety in production at Tashidian Coal Mine. In addition, they offered a reference for green and safe production in Jurassic coal mines of western China. [ABSTRACT FROM AUTHOR]

Based on the field conditions of Huojitu Coal Mine of Shendong coal field in Northern China, eight dynamic numerical models are developed to examine the interaction between the shield and surrounding rock of the longwall panel. The models take into consideration the actual mining process and are developed to simulate the change in the working resistance of the shield for the same shallow burial but with different mining height conditions, ranging from 2.5 to 7 m. The results show that the average shield resistance during the basic roof initial and periodic weighting is 7 to 20 times the weight load of the mining height under the same shallow stratum conditions. By means of dimensional analysis, a relation between mining height and shield resistance in longwall panels is derived theoretically and is fitted by using the shield resistance data under different mining heights monitored in the numerical simulations. This work has led to a new method for the selection of shield support in shallow coal seams. [ABSTRACT FROM AUTHOR]

To study the fracture law of different overlying strata during mining of the lower protective seam under a close distance coal seam at a depth of 1000 m, roof bed‐separation and fissure imaging monitoring were carried out in the inlet roadway ahead of the working face of the JI15‐31040 lower protective seam in the Pingdingshan XII coal mine. The results show that with the overlying strata of the lower protective seam being affected by mining, the displacement of the overlying strata in the roof of 0–4 m was relatively large, and the fissure development was very significant. The displacement of the overlying strata in the roof of 4–12 m was relatively small, the fissures in this range were mostly randomly closed fissures, and this area was stable. Bed separation occurred in the roof of 12–16 m, and the bed‐separating fissures were mostly located at the coal–rock junction. During the mining process, the fractal dimension of the roof gradually increased, and the fissure network developed continuously. The changing process of the fractal dimension of the roof can be roughly divided into three typical change areas: less influence area (distance from the working face > 70 m), rapid increase area (15 m < distance from the working face < 70 m), and stable area (distance from the working face < 15 m), and the area within 70 m in front of the working face was the key supporting area. Under the influence of mining, the roof fissures showed an overall change trend of "from zero to some, from short to long, from small to large, and from narrow to wide." This study can provide a reference for mining of the protective seam in deep coal seams. [ABSTRACT FROM AUTHOR]

Purpose. Coal mining activities by underground mining method will disrupt the equilibrium of the surrounding rock mass. In underground mining, determining the influence zone of exploiting the longwall on the topographic surface is extremely important. Through analyzing the displacement and deformation of the rock mass when exploiting the longwall 21103 in Seam 11 at Nui Beo Coal Mine, the influence is to be determined of exploiting this longwall on the topographic surface in order to forecast the boundary of the influence zone that may affect the surface works, ensuring the safety of such works. Methodology. To achieve the research results, actual field survey methods, data analysis and numerical modeling methods are used in this article. Findings. Analysis of the results from the numerical model determined that the total height of the collapsed and cracked area when exploiting the longwall 21103 in Seam 11 is about 85 m. The subsidence to the ground is about 10 cm from the center of the subsidence area. The analysis results from the numerical model also show that the roof collapse angle is 69°, thereby determining the area of influence on the topographic surface in the range of 55 m. Results on the cross-sections show that the boundary of the influence of mining on the topographic surface is the closest section 6–6, which is about 12 m. The stability time of the collapsed rock mass is determined to be about 5 months. Originality. On the basis of UDEC (Universal Distinct Element Code) software, the authors have developed a simulation model for the mining process of the longwall 21103 in Seam 11. Analysis of the results from the model has shown the state of displacement and deformation of the surrounding rock mass. In this study, the numerical modeling method is applied to simulate the displacement of the longwall, which is consistent with the actual production of the underground mine. This enabled the authors to assess and identify the affected area with the required degree of reliability. Practical value. The research results of the paper are used as a basis for implementation in actual production at Nui Beo Coal Mine. On the basis of the analysis of the surrounding rock mass displacements and deformations, when exploiting the longwall 21103 in Seam 11 at Nui Beo Coal Mine, the affected boundary of the topographic surface has been determined. At the same time, the stability time of the collapsed rock mass is also calculated. So, this is also the basis for Nui Beo Coal Mine to plan construction works on the site outside the determined affected areas to ensure the safety of such works. [ABSTRACT FROM AUTHOR]

Underground resource exploitation has seriously damaged the surface ecological environment and underground water system. As an effective control measure, the filling mining process has greatly reduced the surface subsidence. As a branch of the filling mining process, the continuous mining and continuous backfilling (CMCB) method solves the contradiction of mining and filling in this process. The control index of filling ratio of even number lane (FRE) was presented to investigate the technical advantages of the CMCB method. The numerical analysis model was used to investigate the laws such as deformation characteristics of the surrounding rock, stress distribution, and plastic area distribution characteristics of backfill under four typical cases. As a consequence, the FRE effect law on overburden deformation and the roof control function of the backfill was disclosed, and overburden rock deformation control solutions were provided. According to the results, the overburden deformation varies dramatically when the FRE decreases, and it rises greatly when the even-numbered lane backfill (ELB) is not contacted with the roof. The contacting condition and filling condition of the odd numbered lane backfill (OLB) are connected to the distribution of stress and plastic zone. The backfill transmits the rock beam load by building a composite support system with the roof and floor rock layers, and it accomplishes the backfill's roof control function by combining the primary and secondary load-bearing and synergistic load-bearing connections between the backfills. Measures such as differential FRE, differential strength, non-uniformity of filling lane, and synergistic bearing of temporary support and backfill may help to decrease deformations and internal cracks in the surrounding rock. This measure has been successfully implemented in the field, serving as an experience for the application of the CMCB method. [ABSTRACT FROM AUTHOR]

The room-and-pillar technique is a self-supporting mining method in which part of the ore is left unmined as pillars. To increase the mining efficiency, the pillars are later recovered partially or completely. This research aims at evaluating and comparing four methods of pillar recovery, namely, 'pocket and wing', 'open ending', 'split and fender', and 'Christmas tree' which are of particular importance. The performance of these methods of pillar recovery in the Tabas coal mine was evaluated by numerical modeling. To do this, a series of three-dimensional simulations were carried out in which the pillars were excavated according to the sequences specified by each method of recovery. The behavior of chain pillars, barrier pillars, and the roof was monitored and compared for the different cutting sequences. Using the shear strength reduction technique, the safety factor values were calculated during the different recovery methods until the verge of roof caving. The results revealed that the method of split and fender leads to a softer drop of the safety factor, and consequently, more controlled caving of the roof. [ABSTRACT FROM AUTHOR]

The 110 mining method is a high-efficiency entry-retaining technology without coal pillars or filling materials. At present, there is no precedent for its application in the Huaneng Group. In order to introduce this technology, it is planned to carry out experimental application research in the Yuwang NO.1 Coal Mine of the Huaneng Diandong mining area. The Yuwang NO.1 Coal Mine is a typical coal and gas outburst mine with a coal seam group. In view of the introduction of the 110 mining method under these conditions, first, the geological conditions of the Yuwang NO.1 Coal Mine in the Huaneng Diandong mining area are analyzed, the geological characteristics of the test mining face are summarized, and the practical feasibility of the 110 mining method is analyzed according to the geological characteristics of "one soft, one low, two high, and two complex". Then, according to the engineering experience, calculations, and analysis, the key parameters of roof cutting of the test mining face in the Yuwang NO.1 Coal Mine are obtained, and with the help of a numerical simulation, the roof-cutting height, the roof-cutting angle, and the blasting parameters are numerically simulated and analyzed. The roof-cutting key parameters of the test face are obtained as follows: the roof-cutting depth is 7 m, the roof-cutting angle is 15°, and the blasting method is continuous hole blasting with 500 mm spacing. After that, according to the coal and gas outburst and the occurrence conditions of coal seams in the Yuwang NO.1 Coal Mine, a variety of gob closure design schemes and gob gas drainage design schemes are put forward, and the field effect investigation scheme is given. After the tunnels and open-off cut of the test coal mine's first mining face are connected, under the guidance of the above research results, the field blasting test and the 110 mining method entry-retaining test are further carried out to verify the rationality of the design of the roof-cutting parameters and the feasibility of the gangue wall closure and gas drainage design. Furthermore, in the process of the field practice, continuous research is carried out on the stope pressure law and the adjacent layer gas drainage technology under the 110 mining method, and finally, the 110 mining method application technology system in the Diandong Mining Area is formed. [ABSTRACT FROM AUTHOR]

The disturbance depth of traffic load has a direct impact on the stability of a room-and-pillar mining goaf. To quantitatively calculate the relationship between the traffic load disturbance depth and influencing factors, 49 groups of horizontal combinations of different influencing parameters are designed in this study, based on the orthogonal experimental design method. Midas GTS is used to simulate and obtain the corresponding traffic load disturbance depth data. A multivariate linear regression analysis of the traffic load disturbance depth is conducted, and a regression formula for calculating the traffic load disturbance depth is established. According to the traffic load disturbance depth, goaf depth, and the stability of the roof, coal pillar, and base plate under traffic load conditions, a judgment flow of the room-and-pillar mining goaf treatment method under traffic load conditions is established, and it is applied to the reconstruction and expansion project of the Jixi section of the Dan-A national highway. The results show that a geogrid can be used for treatment purposes when the traffic load disturbance depth is 1.5 times lower than the depth of the room-and-pillar mining goaf, or when the traffic load disturbance depth is 1.5 times greater than the depth of the room-and-pillar mining goaf but the roof, coal pillar, and base plate are stable. Additionally, grouting is needed for treatment in other cases. The results of this study can provide a scientific basis for the selection of treatment methods for room-and-pillar mining goafs underlying highways in the future. The results are of great significance in the field of engineering for the safety measures concerning highway room-and-pillar mining goafs. [ABSTRACT FROM AUTHOR]

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

There are several massive deposits around the world with different geological characteristics. Thus, different mining methods and strategies are applied based on the particularity of each method and mine experience. Particularly, in this work, we review and summarize the underground exploitation of some world-class base metal deposits based on Poland and Chilean experiences. Here, the main geological and mining parameters of Poland and Chilean mines applied in massive deposits are reported and analyzed. In Poland, mainly room and pillar methods (and variants) have been applied in massive deposits. Here, back-filling is required to maintain the mine's stability due to the large deposit size and open areas. In Chile, the block caving method is commonly used in massive underground deposits where less development is required. Here, the cave is naturally filled with broken material and a large subsidence zone is generated. In this review, it has been observed that different underground methods and strategies can be effectively used in massive deposits. Some parameters that influenced the method selection are mainly related to rock mechanics, ore recovery and dilution, subsidence zone, extraction rate, and mining experience. Here, key mining variables and parameters such as productivity, support, and equipment, as well as various issues related to the world-class deposit are studied. Additionally, a comparison between both experiences is presented, highlighting the main geological and mining parameters. This study can be used as a reference to evaluate the different option of underground mining methods to be applied in future massive mine projects with similar geological characteristics. [ABSTRACT FROM AUTHOR]

The more complex the hydrogeological conditions of a mine, the more likely the coal seam is to experience water inrush during the mining process, and the greater the degree of the water inrush hazard. The scientific and reasonable prediction of water inrush in mines with complex hydrogeological conditions is of great significance to the safe and efficient operation of coal mines. Taking the roof water inrush problem of the No. 3 lower coal seam in the Jisan Coal Mine as the research object, the factors affecting the roof water inrush of the coal seam were comprehensively considered from three aspects: the aquifer property, the mining fracture development and the geological structure. The evaluation index system was constructed by selecting 10 factors, including the aquifer depth and thickness, core recovery rate, brittle–plastic rock thickness ratio, number of water-resisting layers, development height of the water-conducting fracture zone, fault density, frequency density, scale index and variation coefficient of the coal seam floor dip angle. At the same time, based on the dual influence of subjective and objective weighting, the scientific and reasonable weight of each factor was ensured. The AHP-CRITIC composite weighting method was used to calculate the comprehensive weight of each factor. Finally, the roof water inrush risk prediction model was constructed. According to the prediction results, the study area was divided into a low-risk area, medium-risk area and high-risk area. Compared with the actual situation, the prediction results were basically consistent with the actual situation, and the prediction results can provide the basis for the prevention and control of water in a coal mine. [ABSTRACT FROM AUTHOR]

In recent years, strong seisms are triggered frequently during deep coal mining process, which seriously threatens the safety of underground miners and ground residents, as well as restricts the productivity and effectiveness of mining activities. It is essential to investigate focal mechanisms and source parameters of seisms for understanding the rock fracturing, the rupture scale and the failure mode. However, the focal mechanism and the internal relationship among source parameters lack systematic research in deep mines. More than 11473 seismic events have been recorded so far in the No.6 mining area of Dongtan Coal Mine, among which 27 events with moment magnitude MW ≥ 1.0 occur in panel 06/63. Results show that Savage model (ω3) has well applicability for spectral analysis in 27 observed seismic events and accords with the characteristics of fast attenuation of displacement amplitude in high frequency. Focal mechanisms of seism events are mainly dominated by pure shear failure and mixed shear failure. Based on the moment tensor inversion (MTI) method and genetic algorithm (GA, a deep learning method), the correlation of source parameters is revealed, and the scale and occurrence of source rupture are calculated effectively. The new preliminary criterion for determining rock fracture type by the dip angle δ and inclination angle θ of fracture surfaces is proposed. Formation mechanisms of strong seisms can be categorized into two types by using the spatial distribution characteristics of fracture surfaces and P–T projection. Meanwhile, we believe that the two mechanisms cross and promote each other under mining disturbance and overburden loads. Highlights: The moment tensor inversion (MTI) method is used to distinguish the focal mechanism of mine seisms. The correlation of source parameters is developed based on Fourier transform (FFT) method and Savage model. A new preliminary criterion is proposed by the dip angle δ and inclination angle θ of fracture surfaces to determine the fracture types of the rock mass. Formation mechanisms of strong seisms are discussed in detail by using the P–T projection and the fracture surface occurrence. [ABSTRACT FROM AUTHOR]

Impoundments of tailings and active surface subsidence zones induced by ore mining were the major environmental problems. The aim of this study was to investigate the technical feasibility of impoundment of tailings using active surface subsidence zones. The self-designed scaled physical model was conducted to investigate the movement laws of tailings under different heights of granular media and tailings, and then drawing ore in these experiments was performed to simulate the mining process. The experimental results showed that the isolated extraction zone and stoping sequence had a primary influence on the doping process of tailings. Besides, when the height of granular media was more than that of the isolated movement zone, the uniform mode of drawing ore could be decreased the amounts of tailings drawn from drawpoints. Meanwhile, the doping process was not impacted by continuously emission of tailings to active surface subsidence zones. Thus, the field test was employed in the Dabeishan Iron Mine and the mining schemes were put forward based on the experimental data. The results showed that the recovery ratio and dilution ratio were 80.9% and 24.3% respectively, and that the scheme established in this paper could be used in practice. [ABSTRACT FROM AUTHOR]

Coal gangue, as a solid waste produced in the coal mining process, can be disposed by being prepared into backfill materials and then filled in underground goafs, thus controlling strata movement and surface subsidence. However, gangue backfill materials are non-continuous; therefore, research into the surface deformation induced by backfill mining should consider the creep compression behavior of gangue backfill materials. The research took a backfill panel in Tangshan Coal Mine (Tangshan City, Hebei Province, China) as the background. In addition, broken coal gangue was collected in the field to prepare specimens of gangue backfill materials, and their creep compression properties were measured. The corresponding constitutive equation of creep compression was then established and embedded in the numerical software, FLAC3D. By building the numerical model for surface deformation induced by backfill mining, the surface deformation above the backfill panel under conditions of different creep durations of backfill materials was simulated and evaluated. In addition, two measuring lines were arranged on the surface to monitor changes in surface subsidence. After surface subsidence stabilized, the maximum surface subsidence was 163.4 mm, which satisfied the fortification criterion of surface buildings. This means the backfill mining did not affect nearby buildings. The results provide a theoretical basis for predicting surface deformation induced by backfill mining and its effective control. [ABSTRACT FROM AUTHOR]

Copyright of Journal of Mine Automation is the property of Industry & Mine Automation Editorial Department and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Backfilling technology is not always used by mining enterprises, which is conditioned by technological and economic factors, such as the need for high mining rates and costs for the technological processes of transporting backfill materials from the daylight surface to the mined-out space. This concerns the underground mining of hard coal, which is a strategic energy resource, in the mines of Ukraine. This paper aims to study the effect of leaving the waste bottom rocks in the mined-out space of the longwall face without their drawing to the earth's surface on the geomechanical state of the rocks surrounding the longwall face. The geomechanical assessment of the stress state of the rock mass surrounding the longwall face, when leaving the waste rocks from the seam bottom rocks in the mined-out space, is performed by the finite element method using the Ansys software package. A geomechanical model has been developed and substantiated, which adequately reflects the mining-geological conditions for seam mining within the extraction site, the actual structure and properties of the coal-bearing rock stratum, the parameters of the longwall face and the modified powered support for the processes of leaving the rocks in the mined-out space. The values and patterns have been determined of the decrease in the stress intensity concentrations in the coal-bearing roof mass in the frontal bearing pressure zone and destressing zone with an increase in the ratio of the rock pack thickness to the extracting seam thickness. The relative indicators of the load on the powered support section and the lowering of its roof have been determined by the ratio of the thickness of the rock pack formed in the mined-out space to the extracting seam thickness. The proposed mining method is of significant commercial and research interest for owners of coal mines developing thin coal seams because environmental costs for placing waste on the surface are reduced, and the energy potential of coal is increased due to the separation of waste rocks from coal in underground conditions. The need for a cycle of beneficiation of mined mass is eliminated and the geomechanical conditions of coal mining processes are improved. [ABSTRACT FROM AUTHOR]

Mining of closely spaced multilayer orebodies brings the problems of significant disturbance between adjacent mining layers and drastic structural changes in surrounding rock, which brings the need for a more effective stope support method. Previous research has made sound analysis on filling or bolt support, but neither of them can solely provide ideal support effects. Thus, a novel bolt-filling support method is proposed by utilizing the synergistic effect of rock bolts (cable bolts) and filling. Numerical simulation and similarity experiments were conducted in this research to analyze the support effect of this method for multilayer ore mining. For numerical simulation, the distinct-element modelling framework PFC2D (Particle Flow Code in 2 Dimensions) was applied for four support scenarios based on the calibration of the microscopic parameters of particles in vanadium shale ores. The numerical simulation results show that the number of fractures decreases from 1311 without support through 652 with 95% filling support to 410 with bolt-filling support, which is resulted from the redistribution of the force chains due to support change. On the other hand, a 300 cm ×180 cm × 40 cm similarity model with a geometry similarity constant of 100 was established based on the 4# rock layer profile of Mount Shangheng. Two parts of similarity experiments were conducted to investigate the strains around the stopes in multi-layer ore mining for three support scenarios. The experiment results prove that the highest strain is in the center of the roof on the upper goaf, and the roof-bolt filling support induces smaller strains than zero support and conventional filling support. Finally, an effective bolt-filling support system has been developed and validated, which can improve the safety and the stability of the roofs and interlayers during the mining process of closely spaced multilayer orebody by reducing the overall load and fractures in surrounding rock. [ABSTRACT FROM AUTHOR]

Copyright of Mining & Metallurgy (10057854) is the property of Beijing Research Institute of Mining & Metallurgy Technology Group and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

With the increasing improvement of national ecological standards, the eco-environmental problems caused by super thick coal seam mining in western China are becoming more and more serious. The failure law of weak overburden stratum is an important factor affecting the safe mining of coal. The failure characteristics of weakly cemented overburden under high-intensity mining in the mining area of western China were studied. For this purpose, a case study was conducted in the 1101 working face of the Baituyao Coal Mine in Ürümqi County. Based on the analysis of geological conditions in the study area, we combined empirical calculations with engineering analogy, physical simulation, and numerical simulation to comprehensively analyze the characteristics of mining-induced overburden failure. The study showed that the overburden in the study area had several unfavorable engineering geological characteristics, including ease of softening in the presence of water. The Middle Jurassic Xishanyao Formation is a directly recharged aquifer with a weak water-retaining property. Overburden failure mainly occurred at the two ends of the open-off cut. During the mining process, vertical fissures and bed-separated fissures were periodically developed and closed, and the fissures were interconnected. The overburden was fractured, and the fractured zone showed a trapezoidal shape, tapering off from bottom to top. The heights of the caving zone and the water-conducting fracture zone were 25 and 280 m, respectively, in the 1101 working face of the Baituyao Coal Mine, and the ratio of fracturing to mining height was 14.0. Due to the weakly cemented overburden and the presence of the Neogene weak aquifer, the risk of water and sand bursts still exists in this working face under high-intensity mining. Our findings shed light on the safe mining and environmental protection of the ground surface in coal mine shafts in western China. [ABSTRACT FROM AUTHOR]

Based on the support vector machine theory, the particle swarm algorithm is used to optimize parameters, combined with an analytic network process in order to feature dimensionality reduction of the original data, and a nonlinear algorithm model combining statistical analysis and machine learning analysis is established. Taking the relevant data of overlying rock from two zones in Yingpanhao Coal Mine as an example, the main factors such as rock tectonic development, coal seam dip, mining height, mining operation method and stope width are ranked according to the weights of their contribution to the development height of the two zones in order to determine the main control factors, such as stope width and mining height. Using MATLAB as the experimental platform, 16 sets of two-dimensional mine data similar to the geology of the study area were divided into training and test sets for prediction and simulation, comparing the optimal solutions of various optimization-seeking algorithms to obtain quantitative prediction results based on the nonlinear algorithm model. The kriging interpolation process was carried out by ArcGIS to realize qualitative visualization, and effective classification is carried out according to the natural breakpoint method to obtain six development height divisions of the water-conducting fracture zone in the study area. The results show that the prediction model of the development height of two zones in the coal roof based on the non-linear algorithm has better accuracy and generalization ability. Predicted by the test set, the model's prediction result MAE is less than 10%, and the accuracy is better than the traditional empirical formula method, which makes up for its lack of rough calculation accuracy and solves the problem of not being able to locate the height of roof crack development at specific borehole locations, which has significance in guiding the prediction and prevention of roof water damage and subsidence disasters. [ABSTRACT FROM AUTHOR]

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

With the massive mining of coal resources, coal mining under special conditions such as "Coal mining under buildings and railways and water-bodies" thick coal seams has became a key concern for relevant scholars. As a mining method to effectively replace "Coal mining under buildings and railroads and water-bodies" thick coal seams, paste backfilling is an important part of the coal green mining system. In this paper, by combining theoretical analysis, numerical simulation, and other means, the migration law of overlying rock in paste backfill mining of continuous mining face in thick coal seam upwards under the village was studied. FLAC3D numerical simulation software is used to analyze the upward separation of thick coal seams. Surface subsidence was simulated at each stage of paste backfill mining in the continuous mining face, and the influence of the backfill elastic modulus on the overburden subsidence, the backfill and coal stress, the range of the overlying plastic zone, and the surface deformation was simulated and analyzed. The elastic modulus of the filling body was finally determined. [ABSTRACT FROM AUTHOR]

Most of the coal mines in Southwest China are located in mountainous areas with high vegetation coverage, and most activities are carried out under the mountains. The deformation monitoring and mechanical behavior analysis of the mining area helps reveal the typical mountain deformation and failure mechanism caused by underground mining activities and reduce the risk of mountain collapse in the mining area. In this manuscript, a research method for mountain stability in mining areas is proposed, which combines InSAR deformation monitoring with numerical analysis. Based on the high-precision deformation information obtained by DS-InSAR and the landslide range, a three-dimensional explicit finite difference numerical analysis method was used to reconstruct the landslide model. According to the layout of the coal mining working face, the variation mechanism of overlying stratum stress and the mountain slip in the coal mining process is inverted, and the mechanism of mountain failure and instability in the mining area is analysed. Based on the sentinel data, the experiment performed time series monitoring and inversion analysis of the mountain collapse in Nayong, Guizhou, China. The results show that mining activities a certain distance from the mountain will affect mountain stability, and there are specific mechanisms. From 2015 to 2017, the stress redistribution of overlying strata above the goaf area resulted in dense longitudinal cracks in the landslide body due to coal mining. The mountain is in a continuous damage state, and the supporting force to prevent collapse continues to decrease, resulting in a gradual decrease in landslide stability. Both the time series DS-InSAR monitoring results and numerical simulation results verify the actual occurrence and development of the on-site subsidence. [ABSTRACT FROM AUTHOR]

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

In view of the problem of taking the loose layer as the uniform cloth load while ignoring thickly sticky loose layered arch (TSLLA) of the traditional mining rock overburden structure analysis, combined with the geological conditions of shallow buried thickly sticky loose layer in Shendong mining area, the mechanical model of the mining TSLLA bearing structure is established, the variation law of arch height under different loose layers is analyzed, and the formation conditions of TSLLA are given. The first theory deduced the dynamic evolution mechanism of TSLLA structure advancing. The dynamic evolution of TSLLA structure in 22207 coalface is simulated. The formula for the limit height of TSLLA is derived. A classification method of mining loose layer structure is proposed. The loose layer structure type is divided into load type, arch beam type damage type, and limit arch height type. On this basis, the mechanical model of "arch beam" is constructed. The influence mechanism of the TSLLA on the surface cracks is revealed. The methods and calculation steps for initial surface cracks are presented. The results are verified by the measured data of surface subsidence of the typical thickly sticky layer in Shendong mining area. [ABSTRACT FROM AUTHOR]

In northern Shaanxi province, the coal seam deposit conditions are generally characterized by large thickness and shallow depth of burial, the phenomenon of coal pillar sloughing is serious in the process of working face retrieval, which brings great potential danger to mine safety production. In this paper, by monitoring the stress and damage of the coal pillar in the field section, the stress changes and damage degree of the working face in different mine stages are determined. A new mechanical model of coal pillar in a thick coal seam is constructed and the stress changes and structural breakage law of each section are obtained, the stress transfers mechanism of overburden structure and the stress damage to coal pillar during the mining process of upper and lower section comprehensive mining working face is analyzed. The stress, microscopic damage, and the degree of sloughing in the coal pillar were significantly reduced by monitoring through the top plate hydraulic cracking pressure relief control program. The results of the study show that the extent of sloughing coal pillar and internal damage is closely related to the length of key block B of the masonry beam formed by the overlying rock layer and the position of the break line, by changing the orientation of the overlying masonry beam rock layer and the position and length of the tendency to break line, the sloughing coal pillar is effectively controlled and its stress and damage are reduced. We have achieved good results in control of the sloughing in the Hanjiawan coal mine and similar mines in northern Shaanxi. [ABSTRACT FROM AUTHOR]

To find out the characteristics of geological structure exposed in the mining process of 12,318 working face in Pansan Mine and grasp its influence law on subsequent coal seams mining, the structure was first determined as the collapse column by means of 3D seismic, transient electromagnetic detection, SYT detection and other methods, and its development characteristics, conductivity and water enrichment were identified. FLAC3D numerical simulation software was used to analyze the characteristics of vertical stress and plastic failure zone in different coal seams during mining. Finally, by comparing the ultimate failure depth of floor and the thickness of waterproof layer in the process of each coal seam directly pushing through the collapse column, the risk of water inrush and the prevention are analyzed. The results show that the exposed geological structure is characterized by weak water-rich collapse column. Under the influence of the mining of the previous coal seam and the activation of the collapse column, the subsequent coal seam is in the low-stress area before mining, which increases the floor failure and causes the activation of the collapse column more easily during mining. [ABSTRACT FROM AUTHOR]

The problems of water scarcity and ecological fragility are common in the loess gully area. To research the distribution and evolution of the overburden fissures and quantitatively analyze them have certain theoretical and engineering significance for realizing the evaluation of overburden damage degree and safe and green mining. This paper takes the 6102 working face of Chuancao Gedan Coal Mine as the engineering background. The development law and distribution characteristics of overburden fissures caused by the mining of shallow coal seams in the loess gully area were studied by the combination of physical similarity simulation, numerical similarity simulation and fractal theory. The results show that the fractal dimension change of the overburden fissures caused by the shallow mining of coal seam groups in the loess gully area can be divided into three stages during the mining process of the working face. Repeated mining causes the activation and development of overburden fissures, the fractal dimension increases significantly, and the regularity of changes weakens. The magnitude of the stress near the working face and the fluctuation times of the stress in the goaf have an influence on the change of the fractal dimension of the overburden fissures. According to the development angle and the fractal dimension of the overburden fissures, the overburden rock above the goaf is divided into the collapse fissure area, the compaction fissure area, and the vertical fissure area. Overburden fissures develop violently in the vertical fissure area, the overburden fissures in the compaction fissure area are mostly transverse fissures, and the overburden fissures in the caving fissure area are irregular. [ABSTRACT FROM AUTHOR]

The rockburst dynamic disasters in the process of deep coal mining become more and more serious. Taking the rockburst occurred in the 23130 working face of Yuejin Coal Mine as the engineering background, we study the characteristics of mining stress field around roadway, the plastic failure morphological characteristics of surrounding rock and the accumulation/release law of elastic energy before and after burst. An analysis model quantitatively describing the physical process of rockburst in the mining roadway is established, and the calculation method of dynamic release of elastic energy in the physical process of rockburst is educed. The mechanism of rockburst in mining roadway is revealed. The results show that an "L-shaped" stress concentration zone is formed within 100 m of the 23130 working face, and the principal stress ratio of the surrounding rock of the transportation roadway is 2.59–4.26. The change of the direction of the maximum principal stress has a significant effect on the burst appearance characteristics. The failure strength of different sections of the mining roadway is characterized by the elastic energy release value. With the increase of the working face distance, the elastic energy released by burst failure and the expansion variation of failure boundary radius show a nonlinear variation law that tends to decrease steadily after sharp fluctuation. The closer to the working face, the higher the burst risk. At a distance of 10 m from the working surface, the maximum principal stress reaches its maximum value. The butterfly-shaped failure system generated by the surrounding rock of the roadway has energy self-sustainability, and the elastic energy released by the sudden expansion of the butterfly leaf is enough to cause a burst damage of 1.9 magnitude. This work could provide theoretical support for the prediction and prevention of rockburst. [ABSTRACT FROM AUTHOR]

The backfill mining method transports treated tailings to the mined-out area, which not only improves the surrounding environment of the mine but also enables the mined-out area to continue mining and production under the support of the filling body. However, with the growth in the depth and scale of mining, ground subsidence, and backfill deformation are becoming increasingly serious problems. As an example, in the Jinchuan mine, a typical multi-stage filling mining mine in China, the deformation law of surface rock mass and backfill are studied through a method combining field monitoring and numerical simulation. The major findings are as follows: (a) A settlement funnel is formed on the ground, and its radius gradually expands with continuous mining and filling. The location of the settlement center moves toward the surface above the footwall of the ore body, and the maximum subsidence reaches 739 mm in 14.5 years. (b) Three-section mining significantly affects the surface deformation, and the single subsidence center on the upper wall develops into the double subsidence center with the mining and filling. When the three-section mining is finished, the maximum value of the surface subsidence reaches about 1.35 m and the mining area is still in a relatively stable state. (c) The whole filling body presents obvious subsidence, with the development of the multi-stage mining and filling. Bed separation phenomena are found between the filling layers, and the closer to the interior, the more obvious it becomes. The backfill's subsidence characteristics are similar to the surface's; that is, both the subsidence amount and speed are higher on the hanging wall than on the footwall. (d) The backfill mainly shrinks inward in the horizontal direction, and the deformation is mainly manifested as an internal uplift and an external subsidence in the vertical direction. The mass instability of the backfill is difficult because of the insufficient deformation space, and the influence of large-scale deformation on the mining and overlying strata needs to be considered, as well as the local deformation near the rock contact zone surrounding the backfill. The results provide technical support for filling mining in the Jinchuan mine and provide a reference for other projects with similar engineering conditions. [ABSTRACT FROM AUTHOR]

Characterizing the permeability evolution and methane release is of great significance for the safe mining of the high gas outburst seams, as well as coal and gas simultaneous extraction. It contributes to reduce methane emissions from coal mining for greenhouse effect control. Theoretical analysis, laboratory testing, and numerical simulation are widely used methods to characterize the permeability and methane release with the treatment process of pressure-relief mining. However, these methods cannot fully reflect the complexity of filed practice. In this study, we report the effectiveness of protective coal seam (PCS) mining and the pressure-relief area in the protected coal seam (PDCS) based on detailed and integrated field measurements in a Chinese coal mine. To the best of our knowledge, it is the first time to measure the permeability coefficient and gas pressure evolution in the PDCS during the process of PCS longwall mining. The evolution of the permeability coefficient in the pressure-relief area during PCS mining can be divided into four stages: slowly decreasing, sharply increasing, gradually decreasing, and basically stable. The maximum permeability coefficient is 322 times of the initial value and stabilized at 100 times after the goaf compacted. The gas pressure evolution in the PDCS indicates that the strike pressure relief angle is 52.2° at the active longwall face zone, and 59.3° at the installation roadway side. The inclined pressure relief angles at the lower and upper sides of the longwall face are 75° and 78.9°, respectively. The residual gas content and gas pressure of the PDCS in the pressure-relief area are reduced to less than 6 m3/t and within 0.4 MPa, respectively. The field measurements further prove that pressure-relief mining can prevent coal and gas outbursts in PDCSs. The field observations in this paper can serve as benchmark evidence for theoretical analysis and numerical simulations, and also provide insights into realizing safety mining in similar conditions. [ABSTRACT FROM AUTHOR]

The static and dynamic inflow of water from the roof aquifer changes as mining progresses. We used a second-order dynamic model to describe the water inflow process. The parameters of the water inflow model were solved using actual drainage from roof aquifers at nine working faces in the Yuanyanghu mining area of the Ningxia Autonomous Region, China, as well as the peak water inflow values, their locations, the equilibrium values of water inflow, and their initial occurrence locations. The parameters of the second-order dynamic model of water inflow were inversely calculated. The peak values of water inflow without drainage were also calculated. The results indicate that pre-drainage of roof water significantly weakens the intensity of water inflow during the mining process, reducing peak values by more than 72%. The characteristics of the water-conducting fractured zones determine the major drainage locations, while the water-rich and water-conductive nature of the direct discharge aquifer affects the water inflow equilibrium values and initial occurrence positions. The results show that the model parameters and characteristic values of water inflow are determined by the hydrogeological nature of the roof strata, water-conducting fractured zone(s), and mining speed. [ABSTRACT FROM AUTHOR]

There is complex air leakage in the mining of shallow buried close distance coal seam group, which affects the generation and migration of CO in the goaf, and easily leads to exceeding safety limits of CO in the return corner of the working face, which threatens the safety of underground production. To examine this problem, taking Lijiahao Coal Mine as an example, this study analyses the generation law of CO gas, the distribution law of overburden fractures, the characteristics of air leakage in the goaf, the sources of CO in the return corner, and the migration and accumulation law of CO in the goaf under multi-source air leakage in the mining of shallow buried close distance coal seam group through experiment tests, numerical simulations, observations and theoretical analyses. The results indicated that there is an exponential growth relationship between the CO generation rate and the coal temperature, and the critical temperature for rapid oxidation of coal samples is between 70 and 80 °C. The 31,115 working face has complicated air leakage from the working face and ground surface and the goaf of this coal seam. The surface air converges to the return corner through the mining fissure of overburden and 2–2 coal goaf, and the air leakage of the working face flows out from the return roadway through the goaf. The gas leakage in the overlying goaf and the oxidation of residual coal are the main sources of CO in the return corner. The CO generated during the coal mining process and the CO generated by the trackless rubber-tired vehicle operation will increase the CO concentration in the return corner to varying degrees. Under the effect of multi-source air leakage, CO from the overlying goaf and the residual coal in the goaf of this coal seam are migrated to the air return side of the goaf, resulting in the accumulation of CO in the return corner, and both of them have a linear positive correlation with the CO concentration in the return corner. The results of the study have scientific guidance for the control of air leakage and the prevention of CO excess in the goaf. [ABSTRACT FROM AUTHOR]

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Coal mining under the aquifer in Northwest China has brought a series of ecological problems, such as the decline of groundwater levels and the death of surface vegetation. The study of the impermeability of rock strata between coal seams and the overlying aquifers is of great significance to solve these problems and realize water-preserving coal mining (WPCM). Based on mining-induced overburden damage and permeability deterioration, the concept of the "three seepage zones" of overburden is proposed, namely the pipe flow zone, water seepage zone and nominal water-seepage-resistance zone (NWSRZ). Meanwhile, the concept of water-seepage-resistance strata (WSRS) is put forward from the aspects of initial permeability, structural strength, swelling and the stratigraphic structure of the overlying strata. AHP-fuzzy comprehensive evaluation (AHPF) is employed to construct a model to evaluate the water-blocking capacity (WBC) of the WSRS. The model includes three secondary factors and nine tertiary indicators, and the weights and membership functions of the indicators are determined. Subsequently, the model is generalized and applied to the Yu-Shen mining area. The membership degrees are spatially visualized by means of thematic maps. The comprehensive evaluation values Φ of the WBCWSRS of 400 boreholes in the mining area under backfill mining, narrow strip mining, slice mining and longwall cave mining are calculated. Then, the Kriging method is employed to plot the zoning maps of Φ under four different mining methods. In view of different grades of WBCWSRS, three corresponding countermeasures, i.e., mining methods optimization, curtain grouting and underground reservoir construction, are put forward. The fluid–solid coupling embedded in FALC3D software is employed to establish a numerical calculation model to simulate the water table fluctuation of the underground aquifer under the four mining methods, and the reliability of the model is verified indirectly. In this paper, mathematical theory is combined with WPCM to develop an evaluation model of WBCWSRS, which provides a reference for the coordinated development of coal extraction and water resource preservation in arid and semi-arid mining areas. [ABSTRACT FROM AUTHOR]