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An attempt to estimate waste rock and soil shear strengths based upon strengths of intact rock cores Williams, D.J.; Taylor, G.; Herbert, D. Oct 31, 2015

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Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 An attempt to estimate waste rock and soil shear strengths based upon strengths of intact rock cores D.J. Williams The University of Queensland, Brisbane, Australia G. Taylor Glencore, McArthur River Mine, Australia D. Herbert Minserve, Brisbane, Australia ABSTRACT In the planning or expansion stages of an open pit mining project, there may be no waste rock available to undertake the testing of its shear strength parameters. Often, all that is available are the results of limited strength tests on intact core specimens recovered from exploration and geotechnical investigation boreholes. Subsequently, the waste rock that is generated from the rock surrounding the ore body can be tested for its shear strength parameters, although such testing is often very limited and may be non-existent. A comparison of the results from the testing of intact rock cores and the subsequent testing of scalped, loose, waste rock specimens and planes of weakness and weathered waste rock can be used to estimate a “waste rock strength classification” for a particular mining project and rock types, which may be used for the preliminary design of the waste rock dumps. Further, an indicative “soil strength classification” may be established to estimate the shear strength parameters of compacted surficial and residual soils at a site, relative to the intact strength of the parent rocks. The paper describes a case study of the testing of intact borehole rock cores and the subsequent testing of loose, scalped waste rock specimens, planes of weakness, weathered waste rock, and compacted surficial and residual soils, from which estimates are made of the waste rock and soil strength classifications. Key Words: compacted soil, shear strength, strength classification, waste rock 1 INTRODUCTION Due to the extreme difficulty and very high cost of testing rock masses on a sufficiently large scale to produce representative strengths, rock excavation design generally makes use of the strength of intact rock cores down-rated using a suitable rock mass classification system (Hoek et al. 1995). Rock mass classification attempts to account for the discontinuities in the rock mass, which generally govern the stability of underground openings and excavations in rock. The waste rock produced from excavations in rock masses is also generally too coarse-grained to be tested in the laboratory. As a result, the shear strength parameters of waste rock must either be based on rock mass classification systems (for input into the updated Hoek-Brown failure criterion, e.g.; Hoek and Brown 1988), or on the laboratory testing of highly scalped or crushed specimens, both of which may yield unrepresentative shear strength parameters. 2 ROCK MASS CLASSIFICATION SYSTEMS The majority of the development of rock mass classification systems was related to civil engineering tunnelling in hard rock in northern Europe, dating back to Ritter (1879) who described an empirical approach for determining support requirements in tunnel design. Terzaghi (1946) introduced rock loads, carried by steel sets, estimated on the basis of a descriptive classification. Lauffer (1958) proposed that the stand-up time for an unsupported span is related to the quality of the rock mass in which the span is excavated. This was extended Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 by Pacher et al. (1974) to form part of the general tunnelling approach known as the New Austrian Tunnelling Method. Deere et al. (1967) introduced the Rock Quality Designation index (RQD) to provide a quantitative estimate of rock mass quality from drill core logs. Multi-parameter rock mass classification schemes were developed from tunnelling case histories, incorporating all components of the geological engineering character of the rock mass (as summarised; e.g., in Hoek et al. 1995). The most widely known and used of these methods are:  Bieniawski (1976, 1989) – Rock Mass Rating (RMR) system, incorporating significant changes over time in the ratings assigned to the different parameters. The RMR system does not include a stress parameter. The RMR value varies between 81-100 for very good quality rock (a down-rating of up to 19% from the intact rock strength) to <20 for very poor quality rock (a down-rating of >80% from the intact rock strength).  Barton et al. (1974) – Tunnelling Quality Index (Q) for determining rock mass characteristics, and tunnel and excavation support requirements. The numerical value of Q varies on a logarithmic scale from 0.001 to a maximum of 1,000, depending on the RQD, joint sets, joint roughness, joint alteration, and water and stress.  Marinos et al. (2005) – Geological Strength Index (GSI) for estimating rock mass strength and deformation modulus of jointed rock masses, for input into numerical analyses. The GSI removes the heavy dependence on RQD inherent in the RMR and Q systems, placing greater emphasis on basic observations of the rock mass and joint characteristics, and was introduced particularly for application to poorer quality rock masses. However, the GSI system is based upon the assumption that the discontinuities are sufficiently “randomly” orientated that the rock mass behaves as an isotropic mass. Hence, that its behaviour is independent of the direction of the applied loads and that there is no dominant discontinuity orientation. 2.1 Application to mining Civil Engineering tunnels are characterised by relatively consistent geometries. Further, civil engineering tunnelling demands high factors of safety. Underground and open pit mines have irregular geometries dictated by the geometry and grade of the ore body, and tolerate lower factors of safety than long-life civil engineering tunnels that expose the public to risk. Hence, rock mass classification systems have been modified for application to the stability design of underground (e.g. Hoek et al. 1995; Milne et al. 1996) and open pit mines (e.g. Bye and Bell 2001). Modifications to the original rock mass classification systems for application to mining include the Modified Rock Mass Rating System (MRMR; Laubscher 1977; as summarised by Bieniawski 1989). The MRMR adjusts the RMR value to account for in situ and induced stresses, stress changes, and the effects of weathering and blasting. 2.2 Hard rock underground mines For deep underground hard rock mines, the rock is generally fresh, discontinuities are generally tight, and groundwater may be restricted to joints, simplifying the application of rock mass classification systems. Further, the generally good rock quality of deep underground mines, results in the modified rock mass classification systems being appropriate. 2.3 Open pit mines, and underground mines in poor quality rock For open pit mines, and underground mines in poor quality rock, the weathering of the rock will vary, discontinuities are more significant and will open on blasting, excavation or caving, and groundwater effects will also be more significant. This makes the GSI a more appropriate system. For relatively homogeneous rock masses, the GSI value varies from about 90 (a down-rating of only about 10% from the intact rock strength) for very good quality joints (very rough, with fresh, unweathered surfaces) to about 10 (a down-rating of about 90% from the intact rock strength) for very poor quality joints (slickensided, with highly weathered surface and soft clay Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 coatings or infilling). For heterogeneous rock masses, the GSI value of very good quality joints reduces to about 70 (a down-rating of about 30% from the intact rock strength). 2.4 Extension to waste rock and soil Extending the use of rock mass classification to the prediction of the shear strength of waste rock and soils derived from open pit mining has rarely been attempted previously. An attempt to generate a “waste rock strength classification” for a particular mining project and rock types, and an indicative “soil strength classification” for the site soils, is the subject of this paper. 3 MCARTHUR RIVER MINE For the purposes of this paper, data from Glencore’s McArthur River Mine (MRM) were used in an attempt to estimate waste rock and soil shear strengths. The available data included the Uniaxial Compressive Strengths (UCS) and Point Load Index (Is50) values of intact rock cores, and the shear strength parameters of loose, scalped waste rock specimens and planes of weakness, and compacted surficial and residual soils. 3.1 Site description Glencore’s McArthur River Mine is located in the Gulf of Carpentaria in the Northern Territory of Australia, 900 km southeast of Darwin (see Figure 1; Glencore 2015). McArthur River Mine commenced underground operations in 1995, and open pit development between 2007 and 2008. McArthur River Mine is the world’s largest bulk producer of zinc in concentrate form, supplying approximately 3% of the world’s total zinc resources used by all types of smelters. Figure 1. Location of McArthur River Mine, Northern Territory, Australia (Glencore 2015). The current North Overburden Emplacement Facility (NOEF) design was approved as part of the MRM Phase 3 Development Project (NT EPA 2014). Following further overburden geochemistry investigations, it has been identified that the previously proposed and approved overburden management infrastructure is not suitable for life-of-mine overburden management at MRM. Therefore redesign of the approved overburden management infrastructure is required. The redesign of the NOEF and other overburden management infrastructure may potentially incorporate an increased footprint, extending beyond that of the current approved design. The previously approved methodologies and design criteria will be revised to address a significant reduction in the available volume of benign non-acid forming (NAF) overburden material and improved understanding of clay characterisation and performance (McArthur River Mining 2011). Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 Figure 2 shows an aerial photograph taken in July 2015 of the main features of MRM, including the current footprint of the NOEF. As approved in the Phase 3 Expansion of MRM, the ultimate open pit is expected to be approximately 420 m deep, 1.7 km long and 1.5 km wide, covering an area of about 200 ha (McArthur River Mining 2011). The open pit produces about 5 Mtpa of ore at an average zinc grade of 10.7% by mass, plus silver and lead. Ultimately, approximately 500 Mt of waste rock will be generated, to be stored largely in surface overburden emplacement facilities, with backfilling of the pit where practicable. Potentially acid generating waste rock will be selectively placed and encapsulated within the overburden emplacement facilities, making use of compacted clay seals and non-acid generating waste rock. The ore is crushed and put through a heavy medium plant, which separates rejects of up to 30% by mass. The ore is then ground in a SAG mill operating in series with two ball mills, prior to flotation and rougher concentration, further grinding, and final flotation. The tailings generated on flotation of the ore are thickened and stored in a series of surface facilities.  Figure 2. Aerial view of McArthur River Mine (source McArthur River Mine). 3.2 Site climate The climate of the McArthur River region is tropical monsoonal, with a pronounced wet season between November or December and March, and generally dry conditions during the remainder of the year (McArthur River Mining 2011). Mean annual rainfall at the mine site is 715 mm, although it can be twice that in a wet year. Mean annual potential evaporation at the mine site is 3,000 mm. Average daily minimum and maximum temperatures at the mine site are 12 to 29oC in June and 25 to 38oC in December. Flooding of the site is an annual risk, due to the McArthur River and diversion channel passing through the site, and other extreme events include cyclones, droughts and fire. 3.3 Site geology The McArthur Basin comprises Carpentarian and Adelaidean rocks extending from the Alligator River in the Northern Territory east to the Queensland border, including most of Arnhem Land and the Gulf of Carpentaria drainage region (McArthur River Mining 2011). The MRM ore body is about 1.5 km long and 1.0 km wide, with an average thickness of 55 m, and occurs at the base of the pyritic shale member, within the Middle Protezoic age McArthur Group. The Open pit Tailings storage facility NOEF Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 shale member comprises a sequence of inter-bedded pyritic bituminous dolomitic siltstones, sedimentary breccias and volcanic tuffs. The ore body has been folded and eroded along its western margin, which is covered with 30 m depth of soil. The western margin of the ore body is sub-vertical with a strike length of 1.0 km and a vertical height of 200 m. The northern margin inter-fingers with sedimentary breccias and the southern margin grades into thinned nodular barren pyritic siltstone. On the eastern margin, the ore body thickens and is folded, with a strike length of over 600 m. The south-eastern corner of the ore body is down faulted 110 m by the north-eastern trending Woyzbun Fault. Typical dry and wet borehole cores for the MRM site are shown in Figures 3, 4 and 5 for shallow, intermediate and deep depths (Klohn Crippen Berger 2014), respectively. Figures 3 to 5 show the dramatic decrease in the degree of weathering and frequency of fractures (increasing RQD) with depth. Core samples selected for UCS testing are indicated with a circle, and were much easier to select at depth.  (a)                (b) Figure 3. Typical shallow borehole core (16.40 to 19.60 m): (a) dry, and (b) wet (Klohn Crippen Berger 2014).  (a)                (b) Figure 4. Typical intermediate borehole core (70.01 to 73.18 m): (a) dry, and (b) wet (Klohn Crippen Berger 2014).  (a)                (b) Figure 5. Typical deep borehole core (124.19 to 127.28 m): (a) dry, and  (b) wet (Klohn Crippen Berger 2014).   Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015  3.4 UNIAXIAL COMPRESSIVE STRENGTHS OF ROCK CORES Samples Uniaxial Compression Strength tests were carried out on NQ-size or HQ-size (47.6 or 63.5 mm diameter) intact rock core samples, with a minimum length of 2.2 times the diameter. The results of UCS testing of MRM borehole rock core samples are plotted in Figures 6 and 7, as a function of rock type and depth, respectively. Figure 6 suggests that the weakest rock types are the intact brecciated Cooley dolomite and upper pyritic shale, which show relatively little variation in strength around a UCS value of about 50 MPa, and have a representative RMR value in the range from about 40 to 50. The intact west fold shale is the strongest and most variable in strength, with a representative RMR value in the range from about 75 to 85. Some of the variability in strength is likely attributed to the inclusion of tuff beds within the unit. The other intact rock types are of intermediate strength and moderate variability, with representative RMR values of the range from about 45 to 60. Figure 6. Uniaxial Compressive Strengths of different waste rock types. Figure 7 shows a very general trend of increasing intact UCS with increasing depth. Figure 8 shows the UCS data, together with valid Is50 data factored by 23 to represent equivalent UCS values, based on the average correlation factor recommended by ISRM (1985). The variation in UCS values with depth is poorly-correlated, while the variation of Is50 with depth is much better-correlated and steeper. Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015  Figure 7. Uniaxial Compressive Strengths of different waste rock types versus depth.  Figure 8. Uniaxial Compressive Strengths and factored Point Load Index values of different waste rock types versus depth. Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015  3.5 SHEAR STRENGTHS OF WASTE ROCK And SOIL SPECIMENS Multi-stage, shear strength tests were carried out in 60 mm and 250 mm direct shear boxes on loose-placed, scalped waste rock specimens (scalped to -4.75 mm or -13.2 mm for testing in the 60 mm or 250 mm shear boxes, respectively), and in 60 mm direct shear boxes on planes of weakness and weathered waste rock scalped to -1.18 mm. The waste rock types tested included upper pyritic shale, black bituminous shale, lower pyritic shale, Cooley dolomite, and west fold shale. The specimens were tested at their as-sampled moisture content in order to represent their field moisture state, and hence were unsaturated. The nominal normal stresses applied were 100 kPa, 400 kPa and 1,000 kPa, representing dump heights of about 6 m, 22 m and 56 m, respectively, which caused compression of the initially loose specimens prior to shearing, and an increase in the degree of saturation as the volume of the air in the voids was reduced. The waste rock direct shear results are plotted in terms of the derived apparent cohesion and friction angle values in Figure 9. Consolidated, undrained shearing, with pore water pressure measurement, multi-stage, triaxial tests were carried out on Silty SAND alluvium, low plasticity Silty CLAY and Tuff/Clay specimens compacted to 95% of their Standard Maximum Dry Density at Optimum Moisture Content. No attempt was made to fully saturate the specimens prior to testing. The nominal cell pressures stresses applied were 100 kPa, 400 kPa and 1,000 kPa. Limited multi-stage, shear strength tests were also carried out in 60 mm direct shear boxes on compacted soil specimens, again with no attempt to fully saturate the specimens prior to testing. The combined soil shear strength results are also plotted in terms of the derived apparent cohesion and friction angle values in Figure 9. Figure 9. Apparent cohesion values and friction angles of different compacted soils and loose waste rock derived from results of direct shear or triaxial testing. Figure 10 shows the shear strength parameters for the loose waste rock and compacted soils tested plotted together for comparison purposes. The direct shear testing of the loose, scalped waste rock specimens and planes of weakness indicated apparent cohesion values of up to Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 800 kPa and friction angles from 15 to 42o, with the friction angle increasing appreciably as the apparent cohesion decreases. For the loose, scalped waste rock specimens and the planes of weakness and weathered waste rock tested, the apparent cohesion values and friction angles were poorly-correlated.  Figure 10. Apparent cohesion values and friction angles of compacted soils and loose waste rock derived from results of direct shear or triaxial testing. The testing of the compacted superficial and residual soil specimens indicated apparent cohesion values of generally less than 100 kPa and friction angles from 12 to 32o, with the friction angle increasing as the apparent cohesion decreases. For the compacted superficial and residual soils tested, the apparent cohesion values and friction angles were reasonably well-correlated. For each of the waste rock and compacted soil specimens tested, it would be very difficult to select representative “design” shear strength parameters. Figure 10 shows that the shear strength parameters of the scalped rock, planes of weakness and weathered rock, which could be taken to represent waste rock in the overburden emplacement facilities, are significantly higher than those of the compacted surficial and residual soils. However, due to the shear strength parameters comprising a combination of apparent cohesion and friction angle, and the large spread of the data, it is difficult to compare the data sets and to fully quantify the differences. 4 WASTE ROCK AND SOIL STRENGTH CLASSIFICATIONS To better enable comparison between the shear strength parameters of the waste rock and compacted soil specimens, it is desirable to reduce the apparent cohesion and friction angle values to a single shear strength value, which can be done using the Mohr-Coulomb shear strength criterion:  = c + .tan (1) Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 where  is the shear strength, c is the apparent cohesion,  is the applied stress (which may be taken as the vertical overburden stress, equal to the unit weight of each material multiplied by the depth), and  is the friction angle. The values of shear strength obtained for the waste rock and compacted soil specimens using Equation (1) may then be compared with the shear strengths of the intact rock cores, which are given by dividing by 2 the UCS values in Figure 7. Figure 11 shows profiles with depth of the UCS/2 values obtained for the intact rock cores, and of the shear strengths of the loose, scalped waste rock and compacted superficial and residual soils. The profiles with depth are plotted to a log10-scale to reduce the scatter in the data and better show the differences between the data for the different materials, for comparison purposes. From Figure 11, it is estimated that the intact rock degrades 16 to 26-fold (20-fold on average) to waste rock in the dump, and the compacted soils are 26 to 45-fold (34-fold on average) weaker than the intact rock. Hence, the estimated average waste rock and soil strength classifications are 20 and 34, respectively, based on the waste rock and soil types, and laboratory test data, obtained for McArthur River Mine materials. Figure 11. Intact rock core UCS/2, and shear strengths of loose, scalped waste rock and compacted superficial and residual soils. 5 CONCLUSION Exploration and geotechnical borehole investigations for open pits mining projects recover rock cores that are typically subjected to Uniaxial Compressive Strength and Point Load Index testing. However, the shear strengths of the resulting waste rock and soils generated on open pit mining cannot be tested during investigations, and are often not subjected to shear strength testing at all. The McArthur River Mine has available data on the strength testing of intact rock cores, and on the shear strength testing of waste rock, which have been used to estimate a “waste rock strength classification” for that mining project and rock types, relative to their intact strength. Further, the results of shear strength testing of compacted soil specimens have been used to estimate an indicative “soil strength classification” for the compacted superficial and residual soils at the site, relative to the intact strength of the parent rocks. The average waste rock and soil strength classifications estimated for the McArthur River Mine materials were 20 Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 and 34, respectively, which could be used for preliminary design purposes for the parent rock types found at this particular mine. The exercise described in this paper is intended only to demonstrate the extension of simple rock mass classification ideas to waste rock and soils derived from the parent rocks at a particular mine site. The results presented for McArthur River Mine are indicative only. It is not intended that the results be used for detailed design, and they should not be extended to other mine sites or to other rock types. It is suggested that similar strength data be collected for other mine sites and other rock types to enable the results presented for McArthur River Mine to be checked, and perhaps extended. It is noted that the McArthur River Mine intact rock, waste rock and compacted soil strength data show very large degrees of scatter, and it is suggested that further research and testing could assist in explaining this variability. 6 ACKNOWLEDGEMENT The management of Glencore’s McArthur River Mine is acknowledged for providing access and granting approval to publish the strength data contained in this paper. 7 REFERENCES Barton, N., Lien, R. and Lunde, J. 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mechanics, 6, pp. 189-236. Bieniawski Z.T. 1976. Rock mass classification in rock engineering. Proceedings of Symposium on Exploration for Rock Engineering, , Johannesburg, South Africa, 1-5 November 1976, 1, pp. 97-106. (Balkema: Cape Town). Bieniawski, Z.T. 1989. Engineering Rock Mass Classifications. (Wiley: New York). Bye, A.R. and Bell, F.G. 2001. Geotechnical applications in open pit mining. Geotechnical and Geological Engineering, 19, pp. 97-117. Deere, D.U., Hendron, A.J., Patton, F.D. and Cording, E.J. 1967. Design of surface and near surface construction in rock. Proceedings of 8th U.S. Symposium on Rock Mechanics – Failure and Breakage of Rock, pp. 237-302. (SME: New York). Golder. 2013. Geotechnical Investigation Report. Pilot Study for Material Classification – NOEF, to Xstrata Zinc, McArthur River Mine. Report No. 137632036-007-R-Rev0, November 2013. Glencore. 2015. McArthur River Mining. Available from <http://www.mcarthurrivermine.com.au/ EN/ABOUTUS/Pages/Location.aspx> [Accessed on 6 August 2015]. Hoek, E. 1994. Strength of rock and rock masses. ISRM News Journal, 2(2), pp. 4-16. Hoek, E. and Brown, E.T. 1988. The Hoek-Brown failure criterion – a 1988 update. Proceedings of 15th Canadian Rock Mechanics Symposium, Toronto, Canada, 3-4 October 1988, pp. 31-38. Hoek, E. and Brown, E.T. 1997. Practical estimates of rock mass strength. International Journal of Rock Mechanics and Mining Sciences, 34(8), pp. 1165-1186. Hoek, E., Kaiser, P.K. and Bawden, W.F. 1995. Support of Underground Excavations in Hard Rock. (A.A. Balkema: Rotterdam). ISRM. 1985. International Society of Rock Mechanics Commission on Testing Methods. Suggested Method for Determining Point Load Strength. International Journal of Rock Mechanics and Mining Sciences & Geomechanics Abstracts, 22(2), pp. 51-60. Klohn Crippen Berger. 2014. Site Investigation Report to McArthur River Mine. Report No. D09814A10, February 2014. Laubscher, D.H. 1977. Geomechanics classification of jointed rock masses – mining applications. Transactions of Institution of Mining and Metallurgy (Section A: Mining Industry), 86, pp. A1-A8. Proceedings Tailings and Mine Waste 2015 Vancouver, BC, October 26 to 28, 2015 Lauffer, H. 1958. Gebirgsklassifizierung für den Stollenbau. Geology Bauwesen, 74(1), pp. 46–51. Marinos, V., Marinos, P. and Hoek, E. 2005. The geological strength index: applications and limitations. Bulletin of Engineering Geology and the Environment, 64, pp. 55-65. McArthur River Mining. 2011. McArthur River Mine Phase 3 Development, Notice of Intent, 11 March 2011. Milne, D., Hadjigeorgiou, J. and Pakalnis, R. 1996. Rock mass characterisation for underground hard rock mines. Tunnelling and Underground Space Technology, 13(4), pp. 383-391. NT EPA. 2014. McArthur River Mine Phase 3 Development – Environmental Impact Statement. Available from <http://www.ntepa.nt.gov.au/__data/assets/pdf_file/0004/118966/ Chapter-4-Project-Description.pdf> [Accessed on 6 August 2015]. Pacher, F., Rabcewicz, L. and Golser, J. 1974. Zum der seitigen Stand der Gebirgsklassifizierung in Stollen-und Tunnelbau. Proceedings of XXII Geomechanics Colloquim, Salzburg, pp. 51-58. Red Earth Engineering. 2014. Geotechnical Data Report. NOEF Dams, to McArthur River Mine. Report No. 30062014_Rev 1, June 2014. Ritter, W. 1879. Die Statik der Tunnelgewölbe. (Springer: Berlin). Terzaghi, K. 1946. Rock defects and loads on tunnel supports. In: Rock Tunnelling with Steel Supports, pp. 15-99 (Commercial Shearing and Stamping Company: Youngstown, Ohio). 

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