Rock Mechanics Feasibility Study

Simon Nickson, Noranda Technology Centre, Pointe Claire, Quebec Dave Ecobichon, Michel Leclerc, Hemlo Gold Mines Inc. Holloway Joint Venture, and ric Ct, Noranda Mines and Exploration Inc., Matagami Division CIM Bulletin Vol. 89, N 998

The Noranda Technology Centre (NTC) has recently used geotechnical data from diamond drill core and underground mapping to develop rock mechanics mine designs for mining feasibility studies. The principal parameters required to characterize a rock mass are a measure of the rock mass quality, the stress conditions in the rock mass, and the geometry of the ore zone. Geo - technical data and empirical design techniques have been combined to develop a methodology for conducting rock mechanics assessments of advanced exploration projects, or for new areas within an operating mine. The basis for rock mechanics mine design within the Noranda Group is the modified NGI (Q) system of rock mass classification (Barton et al., 1974) and the Modified Stability Graph (Potvin et al., 1989). The Q-system of rock mass classification (Barton et al., 1974) describes the rock mass in terms of six parameters as follows,

whereQ = rock mass quality,RQD = rock quality designation,Jn = joint set number,Jr = joint roughness number,Ja = joint alteration number,Jw = joint water reduction factor, andSRF = stress reduction factor.

Mathews et al. (1981) developed an empirical relationship between the stability number N, and the shape factor, S, of a stope surface. The stability number can be evaluated by:N = Q x A x B x C

whereQ s the Q-system rock mass rating with the stress reduction factor set to one;A is the stress factor;B is the rock defect orientation factor; andC is the design surface orientation factor.

The shape factor, or hydraulic radius, is determined by dividing the surface area by the perimeter. Mathews et al. (1981) proposed a design chart that related the stability number to the shape factor, and defined zones of stability, potential instability, and potential cave. The Modified Stability Graph (Potvin et al., 1989) was developed empirically from the analysis of 175 case histories of open stoping situations in Canada. Based on these case studies and the concepts behind the Mathews method, a stable and a caving zone were identified by relating a modified stability number, N, to the hydraulic radius of the surface, as illustrated in Figure 1. Charts for the determination of the A, B and C factors used in the calculation of N are illustrated in Figure 2. A database of 66 additional case histories of cable supported surfaces was used to identify a support required region on the Modified Stability Graph, where the addition of cable support is required to maintain surface stability.

Contents[hide] 1Rock Mass Characterization 1.1Drill Core Geotechnical Data 1.2Geotechnical Mapping 1.3Rock Mass Classification 1.4Orebody Interpretation 2Development of the Geomechanical Model 2.1Data Handling 2.2Mathews Analysis 3Case Histories 3.1The Holloway Project 3.2The Bell Allard Project 4Discussion 5References

Rock Mass CharacterizationRock mass characterization is the first stage in the compilation of a geomechanical model. Geotechnical data from diamond drill core, in conjunction with geotechnical mapping where available, are used to develop a rock mass classification for different rock units.

Drill Core Geotechnical DataGeotechnical core logging is an important part of any drilling program, whether at the exploration stage or during delineation drilling within an existing mine. Core handling and splitting for assaying purposes destroy potential geotechnical data, even if the core is stored for future reference. Geotechnical data are, therefore, best collected at the same time as geological logging. The important geotechnical parameters to collect include: RQD, fracture frequency, the number of joint sets, joint alteration, and intact rock strength.

RQD, or rock quality designation (Deere et al., 1967), is the basic parameter collected during core logging, and forms the basis of many rock mass classification systems. It was developed to measure the percentage of good rock within a borehole. The original concept of RQD was based on NQ size core, and is defined as the percentage of intact core greater than a threshold value of 100 mm, or twice the NQ core diameter. Core from underground diamond drilling is typically smaller than NQ. Because smaller diameter core is expected to be more sensitive to drilling and handling conditions than larger diameter core, the threshold value used in an evaluation of RQD should reflect this increased sensitivity. RQD measurements at NTC use a threshold value that is based on twice the core diameter. A threshold value of 75 mm (3 in.) would be used for BQ size core and 50 mm (2 in.) for AQ size core. Mechanical breaks due to the drilling and handling process are not included in the RQD evaluation.

Fracture frequency is the number of breaks, caused by natural fractures, per unit length of core. This parameter can be a more sensitive measure of rock mass structure than RQD, which is relatively insensitive in good quality rock. For example, even fracture spacings of 200 mm, 150 mm and 110 mm over 1 metre of NQ size core would all result in an RQD of 100% as indicated in Table 1. The value of fracture frequency provides a more sensitive measure. This is significant for hangingwall or back design, where fracture frequency of bedded or foliated units can be a useful predictor of stability.

It is difficult to determine the number of joint sets in a rock mass from drill core. Oriented core data are rarely available for mining projects. The best estimate can be obtained from underground exposures, if available, or from the structural geology interpretation developed for the particular property.

A computerized borehole camera developed at NTC has been used to provide information on joint orientation. The camera head assembly is designed to rotate 360 degrees and has the ability to locate and orient structure within a borehole. The structural data obtained from a borehole camera survey can provide additional delineation of the number of joint sets within the rock mass. The camera is designed to fit in a 76 mm diameter borehole and presently can be used to a depth of 600 m. Rock mass structure is most visible in a diamond drill hole, but can be difficult to observe with the camera in a percussion hole.

Joint surface alteration is an important parameter in the Q-system of rock mass classification. Basic guidelines for estimating joint alteration (Ja) are currently used, based on simple empirical scratch tests conducted on joint surfaces. If the joint surface can be scratched with a knife blade, the value of Ja is 1 to 1.5. If the surface can be scratched with a fingernail, or feels slippery, the value of Ja is 2. Where the altered surface can be dented with a fingernail, or the joint is infilled with gouge, the value of Ja is 4.

Intact rock strength is another parameter in rock mass classification and in several design methods used to dimension stopes and mine pillars. The Unconfined Compressive Strength (UCS) of intact rock is determined in a laboratory test conducted on specially prepared core samples. This standardized test method is the preferred indicator of intact rock strength, as UCS is used directly in design methods. Approximately 10 to 15 samples from each major rock type are sufficient to characterize the intact rock strength for design. Alternatively, the UCS of intact rock can be estimated from point-load testing (Broch and Franklin, 1972) or according to the ISRM Hardness Index (Brown, 1981).

Joint roughness is difficult to obtain from diamond drill core due to limited surface exposure. An estimate of joint roughness is best obtained from underground observation of the rock mass. If underground exposure is not available, a similar rock mass could be used to provide a preliminary assessment.

Geotechnical MappingWhere underground exposures exist, the rock mass properties of joint set number (Jn), joint roughness (Jr), and joint alteration (Ja) can be directly assessed through geotechnical mapping. These parameters can then be used for rock mass classification and subsequent design. Unfortunately, the input parameters for rock mass classification are often qualitatively described and are, therefore, subjective. Past work conducted by the Geomechanics Group at NTC has focussed on developing methods by which rock mass properties can be measured (Milne et al., 1992).

The number of joint sets is readily determined through conventional line mapping. Joint orientation data is collected from each major rock type and used to develop stereonets and a structural signature for the rock mass. Underground access is often limited in the early stages of a project, but structural line mapping is conducted in all of the available exposed rock units.

Joint roughness and joint alteration data are also collected during the line mapping process. Joint roughness is based on measurements of the small-scale and large-scale roughness. Small-scale roughness (Jrr) is obtained by determining the maximum amplitude over a 10 cm profile, and the large-scale roughness (Jrw) from the maximum amplitude over a 1 m profile. These two parameters are easily measured using a 10 cm steel rule and a 1 m folding rule or HI stick. Each ruler is placed along the joint surface, and the maximum amplitude is recorded for both the 10 cm and 1 m profiles. Small-scale amplitudes less than 2.5 mm are assigned a Jrrof 1.0, while amplitudes greater than 2.5 mm are assigned a Jrrof 1.5. Large scale roughness is determined in a similar manner with Jrwvalues of 1.0, 1.5, and 2.0 being assigned respectively to amplitudes less than 10 mm, between 10 mm and 20 mm, and greater than 20 mm. The joint roughness (Jr) for classification is determined by multiplying the Jrrand Jrwterms. This technique was developed from a survey of joint roughness profiles conducted at different mine sites (Milne et al., 1992). Joint alteration is assessed at the same time as joint roughness measurements, using the same empirical scratch tests conducted on drill core.

Where a structural assessment of large amounts of underground exposure is required in a short period of time, one technique is to conduct a drift walk, where less detail is spent on line mapping, but frequent assessments of Q are conducted at regular intervals. This technique requires a good understanding of the mine geology, and is best done using detailed geology and drift survey plans.

Rock Mass ClassificationThe parameters collected from core logging and geotechnical mapping are used to classify the rock mass using the Q-system. The major rock types are normally used as a basis for preliminary classification. Rock mass domains are then defined through compilation of the rock mass parameters, and lumping of units with similar characteristics. Fault and shear zones are generally treated as separate domains.

Orebody InterpretationThe size and shape of an orebody has an important influence on the stability of potential stope surfaces. The orebody interpretation determines the mining block geometry, which is required for preliminary stope design. A preliminary interpretation of the ore zone is, therefore, required to undertake a rock mechanics feasibility study. Revision of the ore zone interpretation may necessitate a review of initial stope design concepts.

Development of the Geomechanical Model

Data HandlingThe drill hole data and orebody interpretation are compiled within the AMINE environment, an AutoCAD based system developed at NTC for computer aided drafting and design of mining projects. A 3D Geological Interpretation and Modelling (3DGIM) module is used to incorporate diamond drill hole survey, assay, geology, and geotechnical information within AMINE. The drill hole traces and associated data are imported through 3DGIM in three-dimensional space. The orebody interpretation, either as section contours or a three-dimensional mesh, is also imported into AMINE. A section cutting routine in AMINE is used to produce drill hole sections with a compilation of geotechnical data, orebody geometry, and rock units, in order to conduct preliminary stability analyses, as illustrated in Figure 3. This data handling procedure provides a flexible way progressively to build the geotechnical model of a deposit as new information becomes available. Additional drill holes and changes to the orebody interpretation can be generated and imported into an existing model. Geotechnical sections can then be quickly updated in order to conduct design work.

Mathews AnalysisThe orebody geometry, geotechnical sections, and other available data from geotechnical mapping, are used to conduct Mathews analyses for potential stope surfaces. For each potential stope surface, a modified stability number, N, is used to predict a stable shape factor from the Modified Stability Graph, as illustrated in Figure 4.

The A factor in the calculation of the modified stability number is determined from the ratio of intact rock strength (UCS) to induced compressive stress parallel to the surface under consideration. UCS results are obtained from laboratory rock testing using samples collected in the core logging process. Numerical modelling of the mining block is conducted to estimate induced stress levels on stope surfaces. Pre-mining stress conditions are initially estimated using stress/depth relationships developed from in situ measurements in the Canadian Shield (Herget, 1988). The model is also used to evaluate preliminary mining sequences. Calibration of the model occurs at a later stage through in situ stress measurements and monitoring of ground conditions with mining.

Case HistoriesThe Holloway ProjectThe Holloway Project is located within the Abitibi greenstone belt, east of Matheson, Ontario. This gold deposit occurs along a geological contact between ultramafic volcanic rock units to the north, and mafic volcanic units to the south. The deposit has a strike length of approximately 800 m, is located at a depth between 150 m and 900 m below surface, and dips irregularly to the south between 0 degrees and 90 degrees.

The Geomechanics Group from NTC conducted a rock mechanics evaluation (Ecobichon et al., 1993) of the Holloway deposit. The objective of the study was to develop an understanding of the rock mass and determine feasible stope dimensions. The rock mass properties were evaluated through intact rock property testing, compilation of drill core geotechnical data, geo - technical mapping, borehole camera surveys, and rock mass classification. The geotechnical properties of RQD, fracture frequency and ISRM (Brown, 1981) rock hardness were recorded by site geological personnel during the core logging process. Compilation of these data indicated that RQD for all rock types was generally greater than 75%, as illustrated in Figure 3, and was, therefore, a relatively insensitive measure of rock mass structure. Fracture frequency was plotted on separate sections, as illustrated in Figure 5, and provided a better indication of rock mass structure. Geotechnical mapping for the purposes of rock mass classification was conducted underground in exploration drifting on the 400 m level. Joint orientation data collected underground indicated that the ore contained two joint sets and an additional random set. The hangingwall and footwall units contained three joint sets each. Borehole camera surveys were conducted in fifteen exploration holes along the strike of the orebody. Discontinuity mapping was conducted within 10 m of significant ore intersections in order to inspect the rock mass adjacent to the orebody. The camera surveys confirmed the existence of a foliation joint set, sub-parallel to the orebody. In general, the foliation and jointing were closed. Foliation adjacent to the ore zone could encourage peeling of hangingwall surfaces, or sliding of footwall surfaces. A low angle joint set was found to be present in all rock units, and may form discrete wedges above underground excavations.

Strength testing of intact rock was conducted through point load testing on diamond drill core at the project site, and Unconfined Compressive Strength (UCS) testing in the laboratory. This testing revealed that the footwall ultramafic rock units to the north of the orebody occur as both weak and strong rocks, depending on their alteration.The footwall ultramafic units were classified using the ISRM Hardness Index as weak (R2). Hangingwall and ore zone rock units were found to be medium (R3) and very strong (R5), respectively.

A Mathews stability analysis was conducted for potential stope surfaces on each geological section, as shown in Figure 4. The N and corresponding shape factor for each surface was determined based on the local rock properties. Stopes were designed to be stable without ground support. The design line on the Modified Stability Graph was taken along the top of the transition zone.Shape factor values were plotted on the longitudinal projection of the Holloway deposit to provide stope design guidelines for different zones of the orebody (Fig. 6). Design shape factors of 6 m for the East Mining Block and 7.5 m for the West Mining Block were selected in order to develop specific stope designs. Table 2 summarizes maximum allowable strike lengths for different stope heights, based on the design shape factors. These dimensions were used at Holloway to develop preliminary longitudinal and transverse stope designs. Some stope surfaces were assessed below the design shape factor and will require additional ground support for stability.

The maximum stable supported width of excavations in ore was established at 8 m. Where the orebody exceeds 8 m in width, a footwall pillar will be required to support the ground above the excavated area. Cable bolt support for ore drifts, using a 2.5 m by 3 m pattern of 3.5 m to 8 m long cables, was recommended. Additional cable bolting will likely be required to support isolated weak footwall surfaces where ultramafic or discontinuous sediments are adjacent to the orebody. These cables should extend 5 m into the footwall on a 2 m by 2 m pattern of birdcage or nutcage type bolts.The preliminary design will be checked and updated as additional information becomes available from diamond drilling and drifting. The design methodology will be repeated in order to optimize the final mine design.

The Bell Allard ProjectThe Bell Allard Project is located within the Matagami mining district in northern Qubec. The deposit is a typical volcanogenic massive sulphide orebody. Massive sulphide lenses are in sharp contact with the overlying rock units, and underlain by a vent structure of hydrothermally altered rock. Vertical faulting is generally associated with these alteration pipe zones. The deposit is located at a depth between 900 m and 1100 m below surface and consists of two irregularly shaped sulphide lenses, with an average dip of 60 degrees to 70 degrees. A rock mechanics assessment (Ecobichon et al., 1994) was conducted as part of the feasibility study for the Bell Allard Project. The objective of the rock mechanics study was to develop an understanding of the rock mass conditions and propose feasible stope dimensions.

RQD and fracture frequency were recorded by project geologists as part of the exploration drilling program. The drill hole traces, in conjunction with RQD and fracture frequency data, were plotted on sections in a similar process to the Holloway Project. UCS testing of diamond drill core samples was conducted at NTC and indicated that the strength of hangingwall, footwall, and ore units were low in relation to the expected stress environment.

No underground access was available at Bell Allard, but geotechnical mapping for the purposes of rock mass classification was conducted in similar rock units at the Isle Dieu mine. An estimate of the number of joint sets, joint roughness, and alteration was extrapolated from the Isle Dieu classification work for the development of Bell Allard stoping dimensions. Preliminary stope dimensions were developed in order to evaluate pillar stability. Numerical modelling of the stope/pillar geometry indicated that stable pillar design would be difficult due to higher induced stresses and the relatively weak massive sulphides. Pillar recovery would present difficult mining conditions and the mine design was, therefore, developed based on mining without pillars. The mine design was further constrained by the expected high stress conditions due to deposit depth and relatively low rock strengths.

The proposed mining approach was to extract transverse stopes in a pyramid sequence on the hangingwall of the deposit. In wide areas, the orebody would be extracted in separate hangingwall and footwall panels. Footwall mining would take place within the stress shadow created by the leading hangingwall pyramid, and result in reduced stress levels around footwall stopes.

Three mining options were developed for Bell Allard:1. 12 m strike length hangingwall stopes with 24 m strike length footwall panels and 34 m sublevel spacing;2. 24 m strike length cable bolted hangingwall stopes with 24 m footwall panels and 34 m sublevel spacing; and3. 12 m strike length full ore width transverse panels with reduced sublevel spacing.

Stope surface dimensioning was developed from the same design methodology followed for the Holloway Project. The hangingwall was recognized to be marginally unstable and was kept small to limit dilution. Option 2 incorporated a larger hangingwall dimension, but included provision for full cable coverage.

DiscussionEvaluation of a mining project requires a basic understanding of the engineering properties of the rock mass, in conjunction with an estimate of the reserves. Compilation of a geomechanical model provides a representation of the rock mass properties and a basis for preliminary mine design.

Characterization of the rock mass is accomplished through a compilation of geotechnical data from diamond drill core, orebody interpretation, and available underground mapping. A program of geotechnical data collection is invaluable in the mine design process. NTC has generally recommended the collection of RQD, fracture frequency, joint alteration, and selected samples for UCS testing from drill core, with an estimate of the number of joint sets and joint roughness. Data handling using combined database and CAD systems, allows for progressive development and updating of the geomechanical model as additional information is obtained. The geomechanical models constructed for Holloway and Bell Allard were part of the feasibility assessment work for each project. The proposed stope dimensions provided mine planning information that was extremely useful in the development of each mine feasibility study. Holloway and Bell Allard project personnel were committed to the collection of geotechnical data during the core logging process, and used this information to develop mine design concepts.

Positive production decisions for both Holloway and Bell Allard have been announced. Validation of the geomechanical models, with additional information from diamond drilling and drifting, is planned in the future to check and update the preliminary mine designs. The design methodology discussed in this paper will be repeated for each deposit in order to optimize final mine designs.

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