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英文原文  A Method for The Design of Longwall Gateroad Roof Support W.Lawrence Geowork Engineering, Emerald, QLD, Australia Abstract:A longwall gateroad roof support design method for roadway development and panel extraction is demonstrated. It is a hybrid numerical and empirical method called gateroad roof support model(GRSM), where specification of roof support comes from charts or equations. GRSM defines suggested roof support densities by linking a rock-mass classification with an index of mining-induced stress, using a large empirical database of Bowen Basin mining experience. Inherent in the development of GRSM is a rock-mass classification scheme applicable to coal measure strata. Coal mine roof rating(CMRR) is an established and robust coal industry standard, while the geological strength index(GSI) may also be used to determine rock-mass geomechanical properties. An elastic three-dimensional numerical model was established to calculate an index of mining induced stress, for both roadway development and longwall retreat.Equations to calculate stress index derived from the numerical modeling have been developed. An industry standed method of quantifying roof support is adopted as a base template (GRSUP). The statistical analyses indicated that an improved quantification of installed support can be gained by simple modifications to the standard formulation of GRSUP. The position of the mathematically determined stable/failed boundary in the design charts can be changed depending on design criteria and specified risk. Key words:Coal mine;Roof control;Support;Design. 1 Introduction Longwall gateroad strata stability is essential to ensure uninterrupted production. In Central Queensland’s Bowen Basin, immediate gateroad roof lithology varies from coal to weak interlaminated material, to strong almost massive sandstone, with localised areas of weak fault affected strata. It is usual for roof conditions within any one mine to vary significantly. Typically, longwall mines in the Bowen Basin have specified gateroad roof support based on past practice. Modifications to gateroad support are generally reactive, due to encountered difficult strata conditions, and less proactive. Current gateroad support design approaches have limitations, which have restricted their applicability and adoption as mine site design tools. A prototype for an improved gateroad support design methodology has been developed that is integrated and systematic, based on rock engineering principals, but requires engineering judgement and experience. There were several broad objec- tives for the design methodology. A consistent and unambiguous definition of strata conditions and behaviour was required. Gateroad roof support needed to be assessed and specified. The method had to provide design calculations and justification for compliance and statutory purposes, and could serve as a framework for a mine strata management system. Mine site support designers must be able to readily use the method to manage uncertainty and risk. The method must be able to be reviewed, modified and expanded. 2 Current Roof Support Design Methods for Longwall Gateroads Numerous roof support design methods have been proposed over the years, but none have gained widespread acceptance by the coal mining industry. There are empirical databases, some proprietary, based on industry practice, which specify gateroad primary and secondary support densities, using a statistical approach. Analytical methods are not appropriate when rock-mass yield due to high mining induced stresses occurs, but may be applicable and adapted to low stress environments. The application of complex post-yield numerical modelling in the design process for excavation support is valid although contentious, and requires a more comprehensive justification and better industry understanding of its strength and limitations. The complete mathematical representation of rock-mass properties and behaviour is a complex issue, which is still outside the capability of current numerical modelling code. Engineers and mathematicians do not have the current capability to fully define rock-mass geomechanical properties and their mathematical representation. Elastic–plastic numerical modelling is a useful tool if used appropriately. It is not exclusively correct or unique, or always superior to other available and accepted design techniques. These aspects have been recognised during recent collaborative Australian Coal Association Research Program research on longwall microseismics, where it was considered that current 3D numerical models lack sufficient validated constitutive relationships, and are forced to make compromises when dealing with complex rock-mass behaviour. Simplified elastic numerical methods have merit and are certainly applicable for more massive sedimentary rock-masses. An assessment of their applicability to weaker, laminated clastic rock-masses is required. Hybrid numerical and empirical methods have been developed for the geotechnical design of undercut and production level drifts of block caving mines. 3 Geotechnical Roof Classification of Longwall Gateroads Two classification schemes were considered appropriate. Firstly, the coal mine roof rating(CMRR), which is an established coal industry standard. Secondly, the Geological Strength Index, GSI with strength parameters included. A recent publication has contended that GSI estimates of rock-mass strength should not be used for coal mine roof problems, where the geometrical scale of the problem is similar to discontinuity spacing. A distinction needs to be made between the GSI classification and the related Hoek–Brown failure criterion. This scale effect and situations where the failure criterion should not be used have been discussed. However, this does not mean that a classification of the rock-mass cannot be made. Indeed, this scale issue is a problem inherent in any rock-mass classification scheme, not just GSI, and for any failure criterion. For example, some mines appropriately use unconfined compressive strength (UCS) as an index or failure criterion, but UCS is also scale dependent and has the same limitations. Within the support design methodology, the rock-mass classification schemes will link mining-induced stresses (or stress index) and required installed roof support. Therefore, the classifications should be independent of environmental and geometrical factors, such as mining induced stresses and excavation orientation and size. A rock-mass classification scheme must also provide rock-mass geomechanical properties to enable the calculation of mining induced stresses. It is anticipated that CMRR will be the principal classification scheme used. However, the single rock-mass classification scheme that is best suited is the GSI derived global rock-mass strength. For numerical or analytical models, Hoek–Brown failure criterion parameters, modulus of deformation and rock-mass strength can be estimated from GSI. Direct utilisation of either CMRR or GSI is included within the design methodology. 4 An index of mining induced stress An index of mining induced stress in the gateroad roof at a location of interest is required. The three-dimensional (3D) stress distribution about a longwall panel including goaf reconsolidation, and the continuous stress redistribution that occurs during panel retreat, is a complex and difficult phenomena to quantify. One approach would be to construct a full elastic–plastic, 3D numerical model. This approach would have limitations to a verified, unique and readily achieved calculation of stress, for several reasons. Generalised model roadway and goaf geometry may not always match the actual geometry. Generalised model roof lithology may not always match the actual lithology and variations. The roof/seam/floor interaction is a complex system and is difficult to model accurately. Rock-mass geomechanical properties, in particular post-yield cannot be fully defined. The geomechanical properties of the goaf, extent and behaviour of strata fracturing and caving, and goaf stress reconsolidation are largely unknown. The model may take many days to complete just a single scenario. While calculated mining induced stress from a detailed elastic–plastic, 3D numerical model may be an appropriate parameter, there is little justification to improved accuracy compared to other methods. An alternative approach is to calculate mining induced stress from elastic 3D numerical models. Calculated mining induced stress in the immediate gateroad roof just outbye of the face-line may not be accurate if rock-mass yield occurs, but as an index of stress, it may be appropriate. An important criterion of its suitability would be how reasonable its relative variation is with changes in input parameters. A significant advantage is that it could be readily calculated for variable scenarios and would be within the range of capability of more geotechnical engineers. Maximum elastic tangential stress in the roof of a modelled gateroad could be considered a better indicator of rock-mass failure than the residual post-yield stress. Undoubtedly, significant rock-mass failure and subsequent stress redistribution do occur, which are not reflected in an elastic model. In the immediate roof of the gateroad, these failures are initiated at a critical mining induced stress. The stress index is a reasonable and appropriate measure of this critical stress, even if it may not agree in absolute magnitude after stress redistribution occurs. For mining induced stresses from an elastic 3D numerical model to be a reasonable representation, several issues influencing the stress distribution must be considered, which include strata fracturing and caving and goaf reconsolidation. For bulking-controlled caving, empirical relationships are used to predict the height of caving (goaf) and fracturing : Hc = Hf = 100h c1h+ c2 100h c4h+ c5 + c3 + c6  (m) (1) (m) (2) Where is the caving(goaf) height above top of extracted horizon, is the thickness of the fractured zone above top of caving zone, h is extraction thickness, and , , , , and are coefficients depending on lithology(Table1). Table1 Coefficients for average height of caving zone[17]. Lithology Compressive strength(MPa) Coefficients c1 c2 (m) c3 (m) c4 c5 (m) c6 (m) Strong and hard >40 2.1 16 2.5 1.2 2 8.9 Medium strong 20-40 4.7 19 2.2 1.6 3.6 5.6 Soft and weak <20 6.2 32 1.5 3.1 5 4 Weathered - 7 63 1.2 5 8 3 Goaf stress–strain behaviour can be been defined (Eq. (3)), based on earlier work, as follows: 1? (ε εm ) σ = E0ε  (MPa) (3) where, and are the vertical goaf strain and stress, respectively, is the initial tangent modulus, and is the maximum possible strain of the bulked goaf material. The initial bulking factor, BF, defines as follows: ε = BF ?1 m BF (4) The initial tangent modulus, , can be defined as a function of the compressive strength of rock pieces,, and the bulking factor, BF: 10.39σ 1.042 E0 = C BF7.7  (MPa) (5) The FLAC3D double-yield constitutive model is used to simulate a strain-stiffening material with irreversible compaction, i.e. volumetric yield, in addition to shear and tensile failure. Upper-bound tangential bulk and shear moduli are specified, with the incremental tangent and shear moduli evolving as plastic volumetric strain takes place. In addition to the shear and tensile strength criteria, a volumetric yield surface or cap has to be defined. The cap surface, defined by the cap pressure, , is related to the plastic volume strain, . The cap pressure, , is not the goaf vertical stress, . The relationship between cap pressure and plastic volume strain is derived from an iterative FLAC2D compression test model, using a one element, 1m1m, grid. Loading was simulated by applying a velocity to the top of the element, which has confined sides and base. The constitutive equation was derived from the iterative results by a Microsoft Excel Solver regression analysis, assuming a linear function. Goaf deformation and material strength parameters are defined as follows(Table2). Table2 FLAC3D goaf reconsolidation parameters. Upper bound tangent modulus 230MPa Poissons ratio 0.3 Density 1.7gm/cc Cohesion 0.001MPa Friction angle 25 Dilation 2 Tensile strength 0MPa ?0.481?1.99e?3σc + 7.83e?3 pc + 0.449BF Table3 FLAC3D numerical model geometrical, geomechanical and geotechnical para- meters. Parameter Range Roadway hight 2-3.4m Roadway width 4.8-6.5m Longwall panel width 200-300m Pillar width 15-45m Depth 60-330m Immediate roof UCS 8-62MPa Ratio of in situ horizontal to vertical stress Rock-mass stiffness dependent. Ranges from 1.2(coal) to 2.0(competent rock) for the major principal stress Rock-mass stiffness E = ?1? D? σci 10(GSI?10) 40 m ? 2 ? 100 ? ? Rock-mass Poisson’s ratio 0.25 for stone,0.30 for coal There are many theories on goaf reconsolidation, based on sound principles. Results from the various formulations do vary significantly. Which, if any, are correct is unknown, as goaf stresses have not been measured. For no other reasons than it is well described, and includes more of the parameters perceived to be important, the goaf stress–strain behaviour as defined Fig. 1. Typical 3D model geometry—horizontal section taken from the top of seam. Fig. 2. Example of FLAC3D model output. is utilised in the calculation of a stress index. The elastic FLAC3D numerical model simulates a single two-heading longwall. Roof and floor strata are composite, uniform continuum. Strong contact is assumed between the coal seam and roof and floor. No discontinuities were modelled. Pillars will always be stable, which means that the actual pillar design must be appropriate and pillars adequately sized for the strata conditions. Some rock-mass geomechanical properties may be derived from the geological strength index. 5 Characterisation of installed roof support A standard measure of the intensity of installed support, widely used within the Industry is GRSUP (ground support rating), given by [4] GRSUP= LbNbCb + LbNtCt 14.6Sbw 14.6Stw (6) Where is the thickness of the bolted horizon defined by roof bolts (m), is the average number of roof-bolts in each bolt row, is the ultimate tensile strength of roof-bolts(kN), is the spacing between roof-bolt rows(m), is the average number of cables in each cable row, is the ultimate tensile strength of cables (kN), is the spacing between cable rows(m), w is the roadway width(m), and 14.6 is a constant that is needed to convert from the original NIOSH equation, which was in Imperial units, to SI units. Roof-bolt variables included in the installed support parameter are length, spacing within and between rows, installed density, and effective density, i.e. Cable variables included in the installed support parameter are ultimate tensile strength, capacity of end anchorage, e.g. barrel and wedge arrangement, row spacing, installed density, installed over conveyor belt structure, grouted, chemically encapsulated or point-anchored, and pretension. The statistical analysis of the database was used to propose modifications to parameters such as N (average number), C (ultimate tensile strength) and S (spacing). 6 Design methodology 6.1 Introduction The design methodology is tailored for roadway development and longwall gateroads. Roof support for bord and pillar first workings can also be assessed. Evaluating roof support using GRSM incorporates several design steps. An initial roof characterisation or classification is required, followed by a calculation of a stress index. Suggested minimum GRSUP is then determined. Finally, primary and secondary roof support patterns are proposed, also considering the influence of factors not assessed by GRSM. 6.2 Rock-mass characterization A classification is required for the immediate 2m of roof, and if a longwall retreat assessment is required, the 4m section above that. Typically, it would be expected that most practitioners would calculate CMRR. Alternatively, the GSI global rock-mass strength may be calculated. Similarly, the intact rock strength should not be overestimated, particularly when using a geophysical correlation. 6.3 Stress index To effectively use GRSM it is important to be able to quickly and accurately calculate a stress index, without having to resort to a FLAC 3D numerical model. Equations to calculate stress index have been developed for two situations; roadway development and longwall retreat. A series of Microsoft Excel Solver analyses were conducted to define equations that could replicate this elastic numerical modelling calculation of stress index. It is recognised that there may be situations where the calculated stress index could be varied. At this stage in the development of GRSM no guidance can be offered about any adjustments. Intersections, both for roadway development and longwall retreat, have different mining-induced stress compared to roadways. Longwall start-up, before regular caving occurs, and major weighting events along the longwall face may have higher abutment stress. As a longwall approaches intersections, there may be an increase in mining-induced stress. Table4 Stress index equation for roadway development—input parameters and constants. Parameters Constants CMRR GSI a -7.66 -7.43 Immediate 2 m roof ( or GSI global rock-mass strength ) 0.033 0.0088 Roadway or excavation height (m) 0.227 0.227 Roadway or excavation width (m) 0.0013 0.0041 Depth of cover (m) 0.00677 0.00681 Solid or rib-to-rib pillar width (m) -0.0013 -0.0013 Ratio of in situ horizontal stress to vertical stress for 0.767 0.772 immediate 2 m roof ;where is the angle between the roadway orientation and 0.280 0.282 the in situ major principal horizontal stress. Only use a positive number between 0and 180 Table5 Stress index equation for longwall retreat—input parameters and constants. Parameters Constants CMRR GSI c -22.40 -28.10 Immediate 2 m roof ( or GSI global rock-mass 0.797 0.168 strength ) Upper 4 m roof ( or GSI global rock-mass -1.064 -0.150 strength ) Roadway or excavation height (m) 0.817 0.794 Roadway or excavation width (m) -0.406 -0.215 Depth of cover (m) 0.0108 0.0101 Solid or rib-to-rib pillar width (m) -0.0129 -0.0098 Longwall panel width,rib-to-rib (m) 0.00021 0.00057 Ratio of in situ horizontal stress to vertical 0.686 0.690 stress for imm. 2 m roof Ratio of in situ horizontal stress to vertical 1.553 0.953 stress for upper 4 m roof ;where is the angle between the gateroad 0.674 0.637 orientation looking inbye and the in situ major principal horizontal stress. Clockwise is positive. The is taken as 20 . For the angle , only use a positive number between 0 and 180 1 1 2 2 1 1 2 2 There are 197 roadway development data points and 78 longwall retreat datapoints. The proposed roadway development stress index calculation is given by Eq.(7), with the input parameters and constants defined in Table4. The proposed longwall retreat stress index calculation is given by Eq.(8), with the input parameters and constants defined in Table5. Both Eqs. (7) and (8) have a correlation coefficient () of 0.99. SIDEV = a+ (bx + b x + ???) + (bx + b x + ???)2  MPa (7) SILR = c+ (d y + d y + ???) + (d y + d y + ???)2  MPa (8) 1 1 2 2 1 1 2 2 where, is the stress index for roadway development, is the stress index for longwall retreat, {,,,,…} are independent input parameters,