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By Guo Wenbing

"INTRODUCTIONThere is a great amount of coal (about 140 billions tons) left unmined under the surface structures, water bodies, railways (referred to as the “three-body”). Coal mining under “threebody” especially that under surface structures has become a major problem in the mining area. The key problem of mining under surface structures is to control overburden strata and surface movements. At present, the major surface subsidence control methods are strip pillar mining, backfilling mining method, room and pillar mining, grouting of bed separations, harmonic mining, and so on. Wongawilli mining method is a high-efficient shortwall mining method developed based on the room and pillar mining, and named after the first successful trial in Australia “Wongawilli” coal seam. The room and pillar mining equipment such as continuous miner is employed in this method. Its major advantage is flexible face layout. It can be used to mine the small wedge coal blocks and those that are difficult to be mined with retreat longwall mining method. It requires less capital investment, short lead time for production, flexible equipment operation and face move and high productivity. It has been applied in some China’s coal mines.The “Wongawilli strip coal mining technology” is a new coal mining method for mining under surface structures. It combines the strip mining longwall layout and Wongawilli high efficient mining technology. This technology can fully utilize the advantages of both the strip pillar and Wongawilli mining methods, and overcomes the disadvantages such as frequent face move, and low mining efficiency in strip pillar mining, and poor ventilation condition and poor long-term stability problems of coal pillars. In this paper, the Wongawilli strip mining method was designed for Wangtaipu mine and its feasibility was studied using numerical modeling and surface subsidence prediction.Geological and mining conditionsAt present, most of the No.15 coal seam reserves totaling 377.07 million tons in Wangtaipu mine are located under surface structures. In order to mine the coal under the surface structures, the Wongawilli strip mining method was employed. An appropriate panel was selected for trial and the surface movement monitoring lines were set up. The trial area is located in panel XV2214 (East) ( re 1). In this area, the floor elevation of coal seam is +634~646m, the surface elevation is +810~817.5m, and the average depth of coal seam is about 174m.The average panel length is 282m and average panel width is 73.5m. The seam dips 1~3° and is 1.8~2.5m thick with an average of 2.15m. The roof and floor rock characteristics are shown in Table 1."

Jan 1, 2015

By Fangtian Wang

The fully mechanized longwall mining was employed in a shallow coal seam that was under the gobs of room and pillar mining and overlain by thin bedrock and thick loose sands. Operational practices have demonstrated that severe accidents would occur including roof structure instability, roof steps, damages of shield supports, and face bumps triggered by the large area roof weighting, resulting in serious threats to the safety of underground miners and equipment. This paper uses Shigetai Mine, located in the Ordos area, Inner Mongolia, as the study site. It fully considers the geological conditions of 3-1-2 coal seam and employs theoretical analysis, 3DEC numerical simulation, and field measurements to analyze the overlying strata movement for the shallow seams under an abandoned room and pillar mine without pillar extraction. The research results serve as an excellent guide to mining in similar geological and mining conditions.

Jan 1, 2014

By Pierre-Yves Declercq, Andre Vervoort

"After the mass closures of entire coal mine districts in Europe at the end of the last century, a new phenomenon of surface movement was observed—an upward movement. Although most surface movement (i.e., subsidence) occurs in the months and years after mining by the longwall method, surface movement still occurs many decades after mining is terminated. After the closure and flooding of underground excavations and surrounding rock, this movement is reversed. This paper focuses on quantifying the upward movement in two neighboring coal mines (Winterslag and Zwartberg, Belgium). The study is based on data from a remote sensing technique: interferometry with synthetic aperture radar (INSAR). The results of the study show that the rate of upward movement in the decade after closure is about 10 mm/year on average. The upward movements are not linked directly to the past exploitation directly underneath a location. The amounts of subsidence at specific locations are linked mainly to their positions relative to an inverse trough shape situated over the entire mined-out areas and their immediate surroundings. Local features, such as geological faults, can have a secondary effect on the local variation of the uplift. The processes of subsidence and uplift are based on completely different mechanisms. Subsidence is initiated by a caving process, while the process of uplift is clearly linked to flooding. INTRODUCTION After the mass closures of entire coal mine districts in Europe at the end of the last century, a new phenomenon of surface movement was observed—upward movement in the area above the past exploitation and in the immediate surroundings. Of course, most surface movement (i.e., subsidence) occurs in the months and years after mining by the longwall method (Peng, 1986). This downward movement continues to occur at a smaller rate many decades after mining has been terminated when the water in the underground excavations is pumped to the surface. This is the case as long as a mine is in operation. However, after the closure and flooding of the underground excavations and the surrounding rock, this movement is reversed, and the surface is uplifted. A total uplift of about half a meter has been recorded to date. This phenomenon has been described in several different coal basins in Europe, for example, in Belgium (Devleeschouwer et al., 2008), France (Samsonov, d’Oreye, and Smets, 2013), Germany (Baglikow, 2011), the Netherlands (Caro Cuenca, Hooper, and Hanssen, 2013), and Poland (Preusse, Kateloe, and Sroka, 2013). To date, research has focused mainly on understanding the phenomenon (e.g., Herrero et al., 2012) and identifying general trends, whereby the link with the rise in water levels is an important issue (Caro Cuenca, Hooper, and Hanssen, 2013). The most accepted explanation is linked to the swelling of clay minerals in the argillaceous rocks in the coal strata (Herrero et al., 2012). Without doubt, the full explanation is complex, and swelling is probably not the only cause of the residual movement. Most coal strata rock is weakened upon contact with water. The increase in water pressure also results in a decrease in the effective stress. Both aspects facilitate the fracturing of rock, which normally would result in further subsidence. However, decreases in the effective stresses also result in the relaxation or expansion of the rock, which induces an upward movement (Fjar et al., 2008). All of these aspects, including the long-term residual subsidence due to compaction under dry conditions, result in a net upward movement. The phenomenon discussed here is clearly different from the upsidence, sometimes observed on other continents (e.g., in Australia, Galvin, 2016). When upsidence occurs, the rock near the surface bends and buckles upwards due to the overstressing of the floors of the valleys."

Jan 1, 2017

By Ry Stone

"There have been many design practices utilised within the coal mining industry to arrive at the minimum densities of primary ground support required during roadway development. This paper demonstrates the practical use of empirical databases, and focuses on the main drivers for ground support as demonstrated in conceptual models.Golder Associates’ empirical databases used for ground support include a Primary Roof Support Database and a Primary Rib Support Database. Both are based on successful ground support designs installed in mines in Australia, the US, the UK, South Africa, New Zealand, and Europe. The term “successful” refers to those designs that were used on a repeated basis for the purpose of roadway development.The Primary Roof Support Database indicates that the major factors influencing successful roof support designs are roof competency, expressed as the Coal Mine Roof Rating (CMRR), and in situ stress. As roof displacement is essentially a function of horizontal stress and roof competency, in order to provide some form of measure as to the likely magnitude of roof deformation during roadway development, the depth of cover is divided by the CMRR to arrive at a “Stress Strength Ratio” (SSR). The database indicates that the higher the SSR, the greater the likelihood that the roof will require higher densities of support.In regard to the Primary Rib Support Database, it is evident from the current database that the primary factors affecting the capacity of rib support required for a successful design are roadway height and depth of cover. In order to provide some form of measure as to the likely magnitude of rib deformation during roadway development, the depth of cover is multiplied by the height of the roadway to determine a “Rib Deformation Rating” (RDR). Similar to the roof, the database indicates that the higher the RDR, the greater the likelihood that the rib will require higher densities of support.These databases have been used to help determine the minimum primary ground support designs required at many mine sites in Australasia, Europe, and the US. This paper will demonstrate the effectiveness and practicality of these databases at two selected mines in Australia and the US.In order to improve the Primary Rib Support Database, this paper will also propose a new rib deformation rating based on the addition of site specific coal strength data for the Australian mines. The proposed rating attempts to capture the main variables that define the behaviour of a buckling column."

Jan 1, 2015

By Dakota Faulkner

Unexpected roof falls often occur in the entries or intersections of active mining sections in underground mines. For large roof falls (greater than 20 ft in height), it becomes dangerous and impractical to clean up the rock debris and re-bolt the newly exposed roof strata. To help mine operators rehabilitate the roof fall areas in a quick, safe, and economic manner, Jennmar Corporation and Keystone Mining Services, LLC (KMS) designed, tested, and developed various impact-resistant (IR) steel set designs, as initially detailed in the proceedings of the 30th International Conference on Ground Control in Mining (ICGCM). Since then, the IR steel sets have been successfully installed in more than 100 roof falls in 43 different coal mines. Significant data has been obtained from these installations, enabling further refinement to and improvement on the design. This paper (1) updates drop test results of the IR lagging panel, (2) presents a case study of the performance of IR arch sets installed 56 months ago at a roof fall rehabilitation site, (3) demonstrates capacity evaluation of the IR steel set, and (4) outlines a preliminary applicability evaluation guideline of the IR steel set for a given roof fall condition. Field evaluation and structural numerical analysis indicate that impact capacity of the system is significantly higher than that of the IR lagging panel, as originally determined. It is concluded that, when evaluating the applicability of an IR steel set, ground control engineers should consider the IR lagging and steel set as an entire system and make a reasonable engineering decision based on actual roof fall geotechnical conditions.

Jan 1, 2014

By Anthony Spearing

There are approximately 68 million rockbolts installed annually in underground coal mines in the USA (Tadolini and Mazzoni, 2006). The most utilized is the passive fully grouted resin rebar bolt, which constitutes almost 90% of the primary support in US underground coal mines. Resin-grouted rockbolts are a very effective and safe support, but their performance still needs to be further optimized and improved. This paper investigates their performance and suggests operating conditions to optimize the performance of resin-grouted rockbolts. In addition, this paper discusses a small-scale test that can also be used to evaluate the performance of rockbolts that is quicker, easier, and less costly than short encapsulation tests.

Jan 1, 2014

By Victor V. Nazimko

Longwall mining technology without interpanel pillars is feasible from a geomechanical point of view. A set of pillarless longwall layouts are proposed including tailentry pillar recovery, headentry pillar recovery and drivage of a new tailentry, partial headentry pillar recovery and combination mining in which advancing panel entries are used for retreating panels. The following mitigative measures are proposed in order to enhance the stability of entries near the panel: driving crosscuts obliquely to the longwall face line, advance support of entries and crosscuts, localized destressing of the chain pillars by horizontal slotting, and partial stowing of the gobs behind the pillars being recovered. A new technology of recovering barrier pillars on the setup entry side is also proposed. This technology uses the same regular longwall face for interpanel pillars extraction.

Jan 1, 1994

By André Vervoort

To learn more about the fundamental roof behaviour in South African coal mines, accurate underground measurements of roof deflection were conducted in bord and pillar panels during development and pillar extraction. To conduct these measurements a survey levelling technique was adapted. The overall system reading accuracy was established as being within 0,4 mm on a 15 m sighting distance. The underground sites were modelled and, for a roadway development, similar symmetrical curvatures were calculated by a linear elastic model, as were measured underground. In a pillar extraction panel, a significant amount of bed separation was measured and the roof underwent an a-symmetric movement due to the close presence of the mined out area at one side of the roadway.

Jan 1, 1992

By David W. Evans

"Induced hydraulic fracture of strata during roof bolt installation is a potentially prevalent, but masked phenomenon within the underground coal industry. Previously reported resin testing programs (McTyer et al., 2014) examined the relationship between resin mixing effectiveness and varying bore hole diameter. The methodology employed within this earlier test program facilitated a further critical area of research – the measurement of back pressures generated within the bore hole during standard rock bolt installation practices. Experimental data has indicated that fluid resin can be pressurised to levels where it exceeds the compressive strength of the strata, inducing hydraulic fracture within the immediate area of the bolting horizon. The routine cycle of roof bolting serves to propagate this effect, progressively fracturing and delaminating the roof during mine advancement. This masked phenomenon can lead to a perception of difficult ground conditions - mining efficiencies and costs are therefore affected, with increased needs for additional support subsequently required to restabilise the inadvertently damaged roof.Further analysis of the parameters associated with resin bolt installation has now been conducted, assisting in the development of an empirical relationship between bore hole pressure, bore hole diameter and bolt insertion times. This relationship has been analysed for 15:1 ratio resins and 2:1 ratioresins, within 28mm and 30mm boreholes. Further to this, load transfer performance has been comparatively assessed for both 28mm and 30mm boreholes, suggesting that for 2:1 resins, acceptable resin mixing and load transfer can be obtained within a 30mm bore hole. The combination of 2:1 resins, utilised within a 30mm bore hole, may well provide the optimal solution to reduce the risk of hydraulic fracture in weaker strata during resin bolt installation.INTRODUCTIONA growing number of industry papers have previously investigated areas of concern associated with the performance of cartridge style resins in roof bolting, predominantly focussing on the effects of plastic film gloving, inadequate resin mixing and the pressurisation of resin within the bore hole. These three effects have a critical influence on the load transfer of the steel bolt element, through the cured resin and into the surrounding strata.Gloving occurs due to the plastic film of the resin cartridge partially encasing or wrapping around the steel roof bolt element – a known phenomenon over many years (Pettibone, 1987). The plastic film creates regions of discontinuity between the bolt, cured resin and borehole, reducing effective load transfer into the strata. Experimentation conducted into the effects of bolt ends fully encased by plastic film (Pastars and MacGregor, 2005) revealed that load transfer can be reduced by 85 to 90% in worstcase occurrences. It is also known that the aggregate filler size used within the resin can assist in shredding and breaking up the plastic film – this effect was observed in a previous study on resin mixing effectiveness (McTyer et al., 2014)."

Jan 1, 2015

By Anton Sroka

The German hard coal production predominantly derives from underground multiseam mining activities. Beside common surface damages, ?inner mine damages? on underground infrastructures (i.e. main roads, cross cuts and shafts) due to mining excavations occur to an increasing extent. As these "inner mine damages" dramatically increased because of the development towards high performance longwall mining technologies, this matter raised to be a severe problem concerning mining economics. To protect durable and long lasting underground infrastructures, precautionary measures such as suitable mine layouts, adapted planning of mining operations and moderate face advances have to be considered. Some examples of "inner mine damages" recorded by means of mine surveying methods are given in this paper. An empirical-mathematical model based on the principles of mining subsidence engineering methods will be introduced. This model allows to minimize ?inner mine damages? and helps to support and optimize mining processes.

Jan 1, 1999

By Houquan Zhang, Yu Wu, Yanlong Chen, Zhenfu Luo, Aihong Lu

"A non-destructive detection scheme for bolt installation quality was designed by adopting the method of non-destructive detection on bolt installation quality. Based on the field measurements, a random non-destructive detection method for determining the characteristic length including anchoring length, free segment length and initial axial load of bolt was discussed, and the characteristics of increasing bolt axial load in the initial stable stage was obtained. The results show that the bolt axial load and deformation of surrounding rocks correlate each other. Compared with the traditional detection technique for bolt installation quality, the non-destructive detection method has obvious advantages such as random, accuracy, and economic. The method provides a reliable basis for bolt support optimization and stability control of surrounding rocks.INTRODUCTIONBolt support in roadway can make the surrounding rocks and bolt form a structure in the early stage of roadway development through the bolt pre-tension, which can maximize the self-bearing capacity of surrounding rocks. Thus it is a form of active support (Kang, 2007). In order to make the bolt play an active supporting role, it needs to ensure the installation quality of bolting. Thus the bolts can supply a high initial anchoring force and control the deformation of surrounding rocks (Shan, 2009; Wei, 2012). Moreover, bolt supporting is a concealed work. Once installation quality problems appear; it will be difficult for effective detection and screening (Xie, 2008).THE PROBLEMS OF BOLT INSTALLATION QUALITY IN COAL MINE AND ITS DETECTIONThe Problems of the In Situ Bolt InstallationDue to the problems in installation environment, machine, and technique, and workers proficiency (Li, 2003), there are many problems exist in bolt installation. If the borehole depth is too long, the anchoring agent will accumulate at the bottom of the bore, which reduces the anchoring length. Furthermore, if the borehole depth is short due to the poor geological condition, the exposed part of the bolt will be too Jong and the pallet cannot cling to the rock face, which prevent the pre-tension from being applied. The pre-tension is very important for bolt installation. If pre-tension is insufficient the rock bolt cannot control the deformation of surrounding rock efficiently and the bolt anchoring force will decrease considerably. The bolt length is another important parameter for anchoring quality, which will affect the control area of bolt supporting. In addition, the cleanness of drill hole has a great influence on the effective anchorage length of bolt. If the drill hole cleaning is not enough, it will directly affect the anchoring agent and the contact area of surrounding rocks, which will decrease the effective anchorage length and initial anchor-force."

Jan 1, 2016

By Yongping Wu

The steeply dipping coal seams in this paper refer to those with the angle of inclination ranging from 35° to 55°. The reserves of the steep coal seams account for approximately 15%~20% of the total coal reserves in China. In the past 12 years, the top-coal caving mining method had been used to mine the steeply dipping thick (11.5-16.4 ft or 3.5-5.0m) coal seams. However, the mining process of top coal caving is complex with low recovery rate and nearly 30% of coal was wasted. Those problems can be solved by using the single pass large mining height (11.5-16.4 ft or 3.5-5.0m) method. But the movement of the overlying strata around the mining face area of large mining height is complex and it is hard to control the stability of ?rock-shield support? system. Therefore, field measurement, numerical simulation and physical (similarity material) simulation methods were used simultaneously to analyze the strata movement, roof structure and rock-shield support interaction characteristics. The results indicate that strata movement, which is the same as mining the steeply dipping seams with conventional mining heights, is asymmetrical about the dipping direction of the face; deformation, failures, and movement of the surrounding rocks are more intense; the roof weighting interval decreases; the weighting intensity increases significantly; and face sloughage occurs frequently. In the gob, the space for sliding and rolling of the roof increases; the broken main roof tends to form an anti-dip pile structure; the backfilling and compactness of broken roof increase at the lower portion of the face, thereby inducing significant unbalance in stress distribution between the upper and lower portions of the face. The overlying strata above the face form a multi-level step structure. The characteristics of contact and loading of roof-shield support system are more complex. The range of shield load variation is larger, and interaction between shield supports occurs frequently. Anti-toppling and anti-sliding devices at the face become more difficult to implement. The research results provide a theoretical basis for the mining practices in the steeply dipping thick seams with a large mining height. The practice in panel 25221 of 2130 coal mine indicates that the accident rates have been greatly reduced and good technical and economic benefits have bee realized.

Jan 1, 2014

By Bruce K. Hebblewhite

BHP Billiton Illawarra Coal operates Dendrobium Mine in an area 10-20km west-northwest of Wollongong in the Southern Coalfield of New South Wales, Australia. The mine recently completed longwall mining in Area 3A adjacent to an overhanging rock feature known as Sandy Creek Waterfall, which had been identified as a significant, and potentially sensitive natural surface feature of the environment. This waterfall lies within the Sydney drinking water catchment area, in a surface terrain of variable topography where non-conventional subsidence behaviour such as valley closure, and far-field horizontal movements are known to occur, over and above conventional subsidence responses to mining. As a part of the conditions of approval to mine, Illawarra Coal undertook to protect the waterfall and the section of Sandy Creek immediately upstream of the waterfall from the effects of longwall mining using an innovative management process and an array of very high resolution monitoring systems. This paper outlines the type of non-conventional subsidence effects that can occur; the range of different monitoring regimes adopted in response to such effects in order to develop an understanding of the valley closure phenomenon in proximity to the waterfall and provide suitable early warning data. The paper also describes the management processes that were adopted, incorporating the use of these monitoring regimes in order to successfully protect the waterfall from the valley closure effects of mining four adjacent longwall panels in close proximity to the waterfall while continuing to maximise recovery of the coal resource in the area. The management structure adopted involved a technical committee, a steering committee, and an external independent reviewer. The technical committee was responsible for design of the monitoring systems, interpretation of the results, and making timely recommendations to the company steering committee who were then able to make informed decisions on when to cease mining each adjacent panel. The steering committee comprised Illawarra Coal management and technical personnel. Although the steering committee took advice from the technical committee, all decisions relating to mining were made by the steering committee. An external reviewer was engaged by the steering committee at the end of each longwall panel to review the results, interpretation and management decisions. This management structure was effective in successfully protecting the very sensitive structure of Sandy Creek Waterfall from potential impacts of nearby mining in the midst of active ongoing natural erosion processes.

Jan 1, 2014

By Daniel W. Su

This paper presents the implementation and evaluation of the hydraulic fracturing technique and Longwall Visual Analysis (LVA) software to mitigate the impact of a 1,000-ft (305-meter) widemassive sandstone channel on a 1,500-ft-wide (457-m-wide) longwall face. Based on a underground roof geology reconnaissance program, four frac holes were drilled and fraced along the center axis of the sandstone channel. To further provide detailed monitoring of the longwall face, the Longwall Visual Analysis (LVA) software was installed to track the face pressure and cavity formation index. In mid-December 2011, the longwall face approached and entered into the western edge of the sandstone channel. The longwall face mined through the center axis of the sandstone channel in mid-January 2012 and advanced outby the eastern edge of the sandstone channel in early February 2012. Results from daily underground observations and production delay analysis confirmed the effectiveness of the hydraulic fracturing program and the validity of the LVA Software.

Jan 1, 2012

By Thomas Rymer

Mathematical modelling of mine subsidence measured at 21 locations in Pennsylvania and northern West Virginia allows for the refinement of current methods of predicting maximum land subsidence associated with longwall coal mining. This new method (RYBAD model) is an extension of that provided by Tandanand and Powell (1984; TP Model). The RYBAD model accomplishes more accurate subsidence predictions (error typically less than 10 percent) for a greater range of mining and geological parameters through the recognition of the following: 1) subsidence increases at a decreasing rate (curvilinear relationship) with greater overburden thickness up to a critical value beyond which subsidence decreases; 2) subsidence is strongly influenced by the stratigraphic proximity of strong rocks to the coal (size of potential caving zone); and 3) subsidence index is related to overburden thickness by a family of curves, each for incremental variations in the effective percentage of strong rock in the overburden.

Jan 1, 1988

By Ashok Kumar, Rajendra Singh, Petr Waclawik

High-grade coal is locked in the shaft protective pillars at 900 m below surface in the Czech part of Upper Silesian Coal Basin. To optimise recovery of coal from the larger size shaft protective pillar, room and pillar (called bord and pillar in India) is trialled to further develop these on smaller pillars. This development followed a constant gallery width of 5.2 m and 3.5 m height. The monitored pillar size in the first developed panel varied from 860 m2 to 1225 m2 of area. The stability of these pillars (i.e., natural supports) at this depth of cover is observed to be a vital issue to control ground movement and safety of the surface structures. A simple calculation of safety factor revealed that application of an empirical method for estimating size and shape of the pillar is inappropriate due to complex geo-mining conditions of the site. Accordingly, geo-technical instruments are installed in the panel for monitoring the stability of the diamond shaped pillar at different stages of the development. Extensive monitoring of the pillar behaviour by these instruments provided an important set of data for such optimum recovery. An established process of simulation-based pillar strength estimation is often used by the mining industry, but calibration is generally done with the help of empirical formulation. However, for the deep-seated pillar, the strength suggested by the existing empirical formulations (except the CMRI formula) is found to be much different from the instrument observations. Using a study conducted at Czech hard coal mines, an attempt is made to develop a strain-softening-based numerical modelling approach to assess pillar performance at deeper cover. The recorded value and nature of the pillar spalling in the field along with the CMRI formula are used to calibrate the developed simulation approach. An assessment of the stability of the developed pillars using simulated models did not indicate any collapse of the natural supports. This paper briefly presents difficulties in pillar design at great depth of cover and field and simulation results of pillar strength estimation in Czech hard coal mines.

Jan 1, 2018

By David A. Newman

In the 1950s, an underground limestone mine in the Camp Nelson Formation was developed from an outcrop exposure. The Camp Nelson Limestone met the requirements for DOT stone and aggregate. The mine ownership changed several times, which resulted in distinct mine designs in Level 1. Based upon the production requirements and an attempt to maximize limestone recovery from a single level, a variety of pillar dimensions and three mining heights (28 feet, 68 feet and 83 feet) were developed. This was sufficient prior to reaching the mineable extent of Level 1. Level 2 introduced previously unknown ground control challenges. These included: ??Pillar spalling and a significant back fall on the Level 1 - Level 2 decline; ??Laminated immediate back and the need to bolt the back on Level 2; ??Delamination of the back adjacent to the Level 1 - Level 2 decline resulting in bolt shearing, and floor heave in isolated areas of Level 2; ??Multiple level interaction, the need to design pillars based upon numerical modeling as columnization between Levels is infeasible: and ??Water inflow as the mine approached river. To meet the challenges, a testing program to determine rock strength and physical properties was implemented along with numerical (LAMODEL and FEM) modeling, and borescope inspections of the immediate back. The efforts to implement rock mechanics were re-focused as Level 2 approached its mineable limit and the initial attempt to drive a decline to Level 3 was unsuccessful. A second decline site was selected based on inter burden stability, a consistent shale parting for the Level 3 back horizon, results from numerical models of the immediate back and multiple level interactions, and the projected low water inflow. The mine is operating on Levels 2 and 3 with long-term planning and ground monitoring for Level 3.

Jan 1, 2014

By Nan Zhou, Jixiong Zhang, Qiang Zhang

"Mining activities under hard roof could easily induce rock burst. Based on the geological conditions of widespread hard roof in China, this paper discusses the type, mechanism of occurrence and prevention of rock burst caused by hard roof from energy evolution point of view. It analyzed the interaction between the solid backfilled body and roof under different backfilling ratios. And by establishing and applying the mechanical model, the deformation of hard roof and energy evolution of the panel under different backfilling ratios were determined, revealing the control effect of solid backfilling on disaster-causing energy. As a result, an innovative method using the solid backfilling method was proposed as a solution to the hard-roof-induced rock burst hazards. Moreover, mine trial was conducted in Panel 6304-1 of Jisan Mine. The result shows that the stress concentration factor of the front abutment stress was only 1.44; the coal dusts in drill holes didn’t exceed the limits and microseism of large energy didn’t occur during mining mining, indicating that solid backfilling mining could be applied as an effective method for preventing rock burst caused by hard roof.INTRODUCTIONSHard roof refers to the thick strata above the coal seam or thin immediate roof. It is strong with poor joint development. After mining, instead of immediate caving, it will overhang for a large area in the gob, resulting in high stress in front of the face and difficulty in gas dispersion. As a result, the face spalls, the roof exposes and the gas accumulates in the gob. With the continuing advance of the face and increase in area of overhang, the stress in the hard roof will eventually exceed its ultimate strength and cause a large area of caving, followed by quick release of energy concentrated in the roof and coal seam, which may cause rock bursts and other dynamic disasters in the mine as well as serious equipment damage and casualties. Therefore, the hard roof becomes one of the main factors causing rock bursts at the face.Hard roofs in China varies in conditions. It ranges from dozens of meters to hundreds of meters in thickness. However, coal reserve under hard roofs accounts for about 1/3 of the total reserve; moreover, nearly 40% of the fully-mechanized coal mining panels are those with hard roofs and more than 50% of the mining areas are associated with problems of hard roofs. Aiming at this problem, researchers at home and abroad have proposed many solutions, e.g. coal pillar support, strength weakening and forced caving, to prevent rock bursts caused by hard roof. These solutions have been employed in some cases with good results. However, due to the variability in geological conditions of mining and hard roofs, the application of the above solutions is limited."

Jan 1, 2015

By Muhammad Zaka Emad, Rangzeb Goraya, Muhammad Usman Khan

"Numerical modeling has been used extensively to assess the stability of surface and underground mines. Tabular, steeply dipping ore bodies are generally mined out by open stoping methods with delayed backfill. Cemented rockfill (CRF) is employed as a backfill material for narrow vein deposits with low production rates. Sublevel stoping methods with delayed backfill are employed to mine out steeply dipping ore bodies. Backfill is exposed to blast-induced vibrations and dynamic loading due to loss of confinement, which could lead to backfill failures. Backfill failures could influence the operating cost of a mining stope and the productivity of the entire mining project. This work shows a detailed overview of the factors contributing to backfill failure. A detailed literature review is presented along with case studies of longhole stoping operations from the Canadian Shield region for this work. The backfilled stopes are analyzed with the finite difference code FLAC3D. The literature review shows that the factors primarily responsible for fill failure are the strength of backfill, stope geometry, mining sequence, and blast-induced vibrations. INTRODUCTION Sublevel stoping methods with delayed backfill are popular in Canadian hardrock and metal mines. Backfill facilitates ground control and greater recovery of ore with pillar mining. It reduces wall slough, thus decreasing ore dilution from adjacent walls. The production of cemented rockfill (CRF) is simple; rock aggregates are mixed together with cement slurry and then placed in a mined-out stope. CRF has a low water content, thus it requires no additional drainage unlike some other backfill. Low water-to-cement ratio enables faster curing rate, hence a faster mining cycle can be achieved (Annor, 1999; Hassani and Archibald, 1998; McKay and Duke, 1989). CRF is generally placed in stopes from the upper sill drift using a truck or a load haul dump machine. Some of the problems associated with CRF are the moderate to high operational cost, the difficulty to get a tight filling, and segregation. Backfill failure is a product of many factors, such as cementation, particle size distribution, placement method employed, water quality, segregation, and blast-induced vibrations. The problem of segregation in aggregates has been observed when aggregate is dumped from a backfill raise (Meech, 2012; Indiana DOT, 2013). Segregation becomes unavoidable when the aggregates are not properly sized (Yu and Counter, 1983; Stone, 2007; Yu, 1989). The exposed backfill zones will contribute to ore dilution due to backfill when subjected to dynamic loading due to fill failures. This paper presents a detailed literature review and numerical modeling results of a mine backfill, along with case studies from longhole stoping operations. The backfilled stopes are analyzed with the finite difference code FLAC3D (Itasca, 2017). The literature review shows that the primary factors responsible for fill failure are the strength of backfill, geometry of stope, sequence of mining, and the vibrations induced by blasting operations."

Jan 1, 2017

By Brent A. Slaker, Michael Murphy, John Ellenberger, Erik Westman

"Photogrammetry, as a tool for monitoring underground mine deformation, is an alternative to traditional point measurement devices, and may be capable of accurate measurements in situations where technologies such as laser scanning are unsuited, undesired or cost-prohibitive. An underground limestone mine in Ohio is used as a test case for monitoring of structurally unstable pillars. Seven pillars were photographed during four visits over a 63- day period. Using photogrammetry, point clouds of the mine geometry were obtained and triangulated surfaces were generated to determine volumes of change over time. Pillar spalling in the range of 0.29–4.03 m3 of rock on individual rib faces was detected. Isolated incidents of rock expansion prior to failure, and the isolated failure of a weak shale band, were also observed. Much of the pillars remained unchanged during the monitoring period, which is indicative of proper alignment in the triangulated surfaces. The photographs of some ribs were of either too poor quality or had insufficient overlap, and were not included. However, photogrammetry was successfully applied to multiple ribs in quantifying the pillar geometry change over time.IntroductionAdequately measuring underground rock mass movements is integral to understanding how rock masses behave and how to interact with them safely and efficiently. Many modern measurement techniques employed in underground mining environments rely on point measurements, such as through extensometers or borehole relief methods (Christiansson, 2006). These techniques, while commonplace, do not provide a comprehensive view of how the rock mass is behaving. The dynamic changes in stress states underground, coupled with the mechanical uncertainty of rock masses, near active excavations, creates a need for measurement systems which better capture the behavior of the rock mass. Photogrammetry is a technology capable of producing the wide-area displacement measurements required for a comprehensive understanding of rock mass movements.Digital photogrammetry is a means of obtaining threedimensional point clouds from digital photographs. By finding the same point across multiple images, all viewing an object from different angles, the location of that point in three-dimensional space can be inferred. Close range digital photogrammetry (CRDP) is photogrammetry applied to measuring objects or scenes less than 100 meters away (Atkinson, 1996), and is used for various functions in underground mining environments. These uses include, but are not limited to, mapping fracture networks (Kim, Kim, & Won, 2005)(Somervuori & Lamberg, 2010), characterizing fractures (Styles et al., 2010) and measuring volumes of blast rock (Heal, Hudyma, & Potvin, 2004)."

Jan 1, 2015