Open Stoping at Golden Grove Under High Stress Conditions

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1 Open Stoping at Golden Grove Under High Stress Conditions Colin Thomson & Ernesto Villaescusa Western Australian School of Mines, Curtin University. ABSTRACT: Mining at Minerals and Metals Group s Gossan Hill mine is fast approaching depths of greater than m and an assessment of stoping practices is required to estimate whether current mining techniques can and should be employed on the orebody below the 124 level. Through an investigation of rock mass characteristics as well as stope analysis using the modified stability graph and back analysis of cavity monitoring surveys (CMS), appropriate stope dimensions and hydraulic radii have been determined. The analysis found stopes would remain stable if designed with hydraulic radii of less than or within the range of 6 to 8, with any increase above these values amplifying instability. Maximum allowable dimensions have been determined for the hangingwall of between 8-25m depending on the number of lifts and rock mass quality. Cablebolting the hangingwall on each sublevel would enable the extraction of stopes with 25m strike spans but should be confirmed with an economic analysis along with the optimal number of lifts. The findings of the research conclude that mining below the 124 level is geotechnically suitable as long as the guidelines described above are adhered to in conjunction with the sustained use of a continuous retreat sequence. INTRODUCTION Golden Grove is a Western Australian mine site located approximately 28 km east of Geraldton and 5 km south of Yalgoo in the Mid West region. The operation there consists of two separate underground mines, namely Scuddles and Gossan Hill owned by Minerals and Metals Group (MMG). The Gossan Hill underground mine, on which this project will focus, is owner-operated by MMG. Mining at Gossan Hill is carried out via sublevel open stoping across several key mineralisations, stretching from near surface to beyond m in depth. The Hougomont orebody is actively being mined from the 544 level down to the 124 level (9828RL-9352RL, m below surface [1369RL]) and is the mineralisation this research concentrates on. A continuous retreat sequencing method is currently in use in the lower Hougomont region, retreating to the North. As mining has progressed into the lower levels and as more ore is discovered at ever increasing depths, stress related issues have started to arise that could potentially threaten both the safety and production of the operation. Mining in the 994 level (938RL, 99 m depth) of the Hougomont mineralisation has been subject to minor wall deterioration owing to the high stress levels encountered, resulting in excess dilution and reduced draw times. Squeezing of drill holes was also encountered and led to the change from a primary, secondary sequencing method to a continuous retreat. As the in-situ stresses are expected to increase with depth, previous mining in similar conditions must be analysed to ensure the ore below the 124 level can be extracted safely and with minimised stress induced consequences. Analysis of previous stoping in the lower Hougomont region ( RL, m depth) has been carried out, including 11 extracted stopes of similar stratigraphic position and applied stresses. Analysis of these stopes should provide an estimate of future stope performance. Ultimately through an assessment of the geology, rock mass characteristics, the geotechnical model and an analysis of previous stoping, an appraisal can be made of the geotechnical suitability of mining the ore below the 124 level (9352RL, 12 m depth).

2 METHODOLOGY A simple methodology was followed to produce a result, which included the following tasks: 1) Assessment of current mining methods and conditions, 2) Determination of site geological conditions, 3) Classification of rock mass and building of a geotechnical model, 4) Validation of uniaxial compressive strength (UCS) data via point load testing, 5) Stope stability and back analysis to determine the affect of stope geometry on stability, and 6) Assessment of geotechnical suitability of mining for the ore below the 124 level. GEOLOGICAL SETTING Both the Gossan Hill and Scuddles deposits are volcanogenic massive sulphides hosted within the Warriedar Fold Belt, part of the Yalgoo-Singleton greenstone belt within the Southern Murchison Province of the Archaen Yilgarn block (MMG, 21). Dacitic and rhyodacitic volcanics of the Scuddles Formation are the main rock types of the hangingwall and bedded tuffaceous volcaniclastic rocks of the Golden Grove Formation make up the footwall (Smith, 23). The Gossan Hill mineralisations are steeply dipping to the west, hosted in a horizon of thinly bedded chert and tuff. The deposit consists of a number of lenses of Zn and Cu mineralisation, extending over a strike of 4 m and a width of 2 m. The Cu occurs in magnetite rich epiclastic rocks stratigraphically below the Zn ore (Smith, 23). Deposits are commonly associated with footwall stockwork (stringer sulphides containing Cu) and chalcopyrite/pyrite mineralisations overlain by banded massive sulphides (most often Zn bearing massive sulphides). The primary rock units found are dacite, dolerite, massive sulphides, chloritic sediments, sediments, rhyodacite and rhyolite (Louchnikov, 211). Locally in the lower Hougomont region, the orebody is steeply dipping to the west (>8 ) with a thickness ranging between 1.8 and 14 m. The footwall consists mostly of sediments classified in the GG6 stratigraphic unit while the hangingwall is mainly formed from rhyodacite of the SC2 unit (Louchnikov, 211). Volcanic intrusives are present in the area with a dacite intrusive bounding the orebody in the South and a dolerite intrusive cutting across the orebody. For the purposes of this research, the dolerite intrusive has been assumed as beyond the scope of the project and hence the footwall has been assumed as totally made up of GG6 sediments and the hangingwall made up of rhyodacite. ROCK MASS CHARACTERISATION A characterisation of the rock mass has been carried out to determine the in-situ properties of the rock mass. Determination of geological discontinuities has been performed as well as an estimation of the intact rock strength and the in-situ stress regime. Geological Discontinuities Large scale discontinuities are present across the mine, however, none have been discovered within a close enough proximity of the lower Hougomont mineralisation to have an effect on the behaviour of the rock mass. Small scale discontinuity data collected using scanline mapping techniques is available, and was collected during the early stages of mining. As increasing amounts of shotcrete are used, the amount of discontinuity mapping has decreased, but where it has been possible to perform it has generally confirmed the historic data (Saunders, 28). Four main joint sets have been identified across the mine site including one foliation, two steeply dipping joint sets and one sub-horizontal joint set. Specific joint set data for the rhyodacite and footwall sediment rock units were collected as well and used in later analysis. Intact Rock Properties Intact rock property data has been collected at Gossan Hill since 1997 and includes a database of over 213 samples. UCS values obtained from the mine site were validated against point load testing conducted on 125 samples collected from Golden Grove. Samples were collected from ten diamond drill holes from below the 124 level. Within each individual drill hole, samples were taken to represent the hangingwall, footwall and orebody. Figure 1 shows the UCS values taken from the mine site database compared alongside the estimated point load testing results.

3 UCS (MPa) In-Situ Stress Figure 1: UCS Data Validation In-situ stress measurements have been recorded at Gossan Hill using the CSIRO HI Cell overcoring method. Five stress measurements have been taken at Gossan Hill although one has been disregarded due to the proximity of the sample area to the Catalpa Fault. Based on this data, as well as raisebore breakout observations, stress gradients and orientations have been determined. Table 1 shows the formula produced from linear regression with depth below surface (D), as well as the trend and plunge for each principal stress. Principal stress GG6 (FW) Table 1: In-Situ Stresses Stress gradient (MPa) Trend ( ) Plunge ( ) 1.7D D D GEOTECHNICAL MODEL Rock Mass Classification Mine UCS Values Tested UCS Values Rhyodacite (HW) Massive Sphalerite (Ore) The Rock Quality Designation index or RQD provides a quantitative estimate of rock mass quality attained from logging drill core (Hoek, 27). MMG has collected RQD values from all diamond drill holes. Values from the same ten drill holes used in the validation of the UCS data have been selected to represent the footwall, orebody and hangingwalls of the area below the 124 level. Table 2 shows the values for each zone. Table 2: RQD Values Zone Footwall Orebody Hangingwall RQD Range The NGI-Q (1974) system of rock mass classification was developed based on numerous case studies to effectively assess rock mass characterisation and tunnel support requirements. It characterises the rock based on an estimated block size (RQD/Jn) and the shear strength between blocks estimated as Jr/Ja (Barton et al, 1974) as well as the stress factors Jw/SRF. For the purposes of the modified stability graph discussed later, the modified Q must be calculated rather than Q and is given in Equation 1. Where: Q = RQD J n RQD = Rock Quality Designation, Jn = Joint set number, Jr = Joint roughness number, and Ja = Joint alteration number. J r J a (1) Table 3 gives the values for Q calculated for the footwall, hangingwall and orebody. Table 3: Q' Values Footwall Orebody Hangingwall RQD Jn Jr Ja Q Rock Mass Strength From a review of the literature, the Hoek-Brown (1997) criterion for estimating the rock mass strength has been determined as suitable method to use. The intact rock properties are determined using Hoek and Brown s formulas before being equated to the linear Mohr-Coulomb envelope to determine the compressive and tensile strengths of the rock mass as well as the rock mass modulus. The Hoek-Brown criterion is given in Equation 2.

4 Modified Stability Number (N') σ 1 = σ σ a 3 + σ ci (m 3 b + s) σ ci The Hoek-Brown constant mi has been estimated from the rock geology and the tables provided by Hoek and Brown (1997). Table 4 shows the estimated mi values as well as the results of the Hoek-Brown methodology. Table 4: Rock Mass Strength Results Footwall Orebody Hangingwall m i σ cm (MPa) σ tm (MPa) E (MPa) (-3.6) -.4-(-3) -.6-(-2) Q Modified Stability Graph Based on a review of the literature, Nickson s research into the uses of the stability graph, published in 1992, was selected as a pertinent approach to follow. The stability graph method (2) attempts to relate the size of an excavation surface to a calculated stability number N. As Nickson s methodology is based on Potvin s (1988) modified stability graph, it is the modified stability number (N ) that is calculated for each surface. N is given in Equation 3. Where: Q = Modified Q Value, A = Stress factor, N = Q A B C (3) B = Rock defect orientation factor, and C = Design surface orientation factor. Factors A, B and C were calculated using the in-situ stress and discontinuity data discussed earlier. A minimum and maximum N value was calculated using the Q range shown in Table 3, and the average of these two values was plotted on Figure 2 against the hydraulic radius (HR) of a stope surface, defined as the area divided by the perimeter. The resulting depth of failure (m) is also shown for each stope surface m 1-2m 2-3m 3-4m 4-5m >5m Hydraulic Radius (HR) Figure 2: Modified Stability Graph Golden Grove Mine

5 Maximum Allowable Width (m) Maximum Allowable Strike Span (m) STOPE STABILITY ANALYSIS In his research, Nickson (1992) applied a mathematical formula to the transition zone developed by Potvin (1988). A simple manipulation of Nickson s formula produces the transition envelope shown in Figure 2. As this formula produces an estimate of stable HR, it can be used to determine maximum allowable spans. This is achieved by rearranging the formula for HR (area over perimeter), then substituting in Nickson s transition line for HR, while holding one dimension fixed. In this case the fixed dimension is the sublevel interval of 3m. Maximum allowable spans have been determined for the hangingwalls, footwalls, North and South walls as well as the stope crowns Footwall Hangingwall Footwall Hangingwall Footwall Hangingwall Footwall Hangingwall Single Lift Double Lift Triple Lift Quad Lift Figure 2: Footwall and Hangingwall Maximum Allowable Strike Span North Wall South Wall North Wall South Wall North Wall South Wall North Wall South Wall Single Lift Double Lift Triple Lift Quad Lift Figure 3: North wall and South wall Maximum Allowable Stope Widths

6 Depth of Failure (m) The results of the footwall, hangingwall and North wall, South wall have been displayed in Figures 3 and 4. The results of the crown allowable spans have not been presented, as there were no restrictions on span length. Calculations have been performed for single and multiple lift stopes of up to 12m in height. Figure 3 clearly shows that the hangingwall is the controlling surface of the stope, as the maximum allowable span ranges between 8 and 25m, whereas the allowable strike span for the footwall is between 25 and over m. For the North and South walls it would appear that the South wall is the controlling surface. However, South wall failure can be ignored as the orebody is being retreated to the North and the previously backfilled stope forms the South wall and to date, there has been minimal failure along this contact. Hence the allowable stope widths for the North wall are between 12 and over m, which is in line with the orebody width of m. This would suggest that any failure associated with the North wall is not resulting from the affects of stope geometry but may be due to other factors associated with ground behaviour such as the induced stress system caused by the continuous retreat method. BACK ANALYSIS OF STOPE PERFORMANCE A comparative analysis was performed between the collected Cavity Monitoring Survey files and the design files in the Vulcan software program. Crosssections of each surface were taken and the maximum depth of overbreak measured. Overbreak is defined as any materials from outside the designed stope shape that rills into the stope. The depth of this failure has been measured perpendicular to each stope surface, as per the method described by Villaescusa (24). This process was performed on 11 previously mined stopes in the lower Hougomont region for each of the five stope surfaces. Figure 5 shows the depth of failure graphed against the hydraulic radius of the hangingwall. From the graph the critical HR can be determined from the trendlines. In this instance the HRcritical is approximately 7. For the analysis of the footwall, the HRcritical was approximately 8. Results from the North wall showed minimal correlation between an increase in the HR and an increase in the depth of failure. As mentioned in the stability analysis, this may suggest that the cause of failure is not associated with stope geometry and may be due to factors such as induced stresses. There was insufficient failure data for the South wall and the crowns to determine a critical HR. The lack of South wall failures is due to the sequencing method whereas crowns have been found to be almost wholly stable Figure 4: Hangingwall Depth of Failure vs. Hydraulic Radius CONCLUSIONS AND RECOMMENDATIONS Conclusions HR critical HR Depth of Failure -1m 1-2m 3-4m 4-5m >5m Based on the collected geological and rock mass data, a basic geotechnical database has been built and used to assess the geotechnical suitability of mining the orebody below the 124 level. Using the empirical modified stability graph method, the maximum allowable stope spans along strike have been determined as 8-25 m, based on hangingwall stability and dependent on the number of lifts and quality of ground. North wall allowable dimensions range from between 12 and > m, again dependent on number of lifts and quality of ground. As these widths are concordant with the orebody widths, resulting failure can be assumed to not be a result of stope geometry but other factors such as induced stresses. Stoping back analysis indicated that a hydraulic radius of up to 6-8 showed minimal overbreak and that increasing this value would amplify instability. From the established stope geometry guidelines, stopes should be designed with strike spans of approximately 25 m with cablebolts installed at each sublevel in the hangingwall to maximize extraction and maintain stability. The continuous retreat method should continue to be used. Ultimately, the orebody below the 124 level is geotechnically suitable for mining given the above constraints.

7 Recommendations for further study As mining progresses deeper and the affects of stress intensify, knowledge of the in-situ stress regime becomes paramount. Additional site testing should be conducted to improve this database to understand and predict the effects of stress at depth. As mentioned previously, there are numerous other factors affecting ground behaviour other than stope geometry. Further study should be conducted into quantifying the affects of these factors such as drill and blast practices, groundwater, time dependent behavior of the rock mass and induced stresses. Economic analysis is also required to determine the affect of overbreak against stope size. This process will also aid in the assessment of the optimal number of lifts and help confirm the recommended strike span lengths. Potvin, Y, Empirical open stope design in Canada, PhD thesis, University of British Columbia, Vancouver. Saunders, P (in prep). A mining sequence analysis and failure criteria application at Gossan Hill mine. Master of Mining Geomechanics, Curtin University. Smith, R E, CRC LEME, 23. Gossan hill cu-zn-au deposit, golden grove, western Australia [online]. Available from < crcleme.org.au > [Accessed: 5 October 211]. Villaescusa, E, 24. Quantifying open stope performance, in Proceedings of MassMin 24, Santiago, Chile (ed: A Karzulovic and M A Alfaro) pp (Mineria Chilena). ACKNOWLEDGEMENTS The authors would like to acknowledge the MMG staff at the Golden Grove site, in particular Wayne Ghavalas, Soma Uggalla and Adam O Hare. REFERENCES Barton, N R, Lien, R and Lunde, J, Engineering classification of rock masses for the design of tunnel support, Rock Mechanics, 6(4), pp Deere, D U, Hendron, A J, Patton, F D and Cording, E J, Design of surface and surface construction in rock, in Proceedings 8th US Symposium on Rock Mechanics, Minneapolis, USA (ed. C. Fairhurst) pp (American Institute of Mining, Metallurgy and Petroleum Engineers: New York). Hoek, E, 27. Practical Rock Engineering [online], Available from: < [Accessed: 5 October 211]. Hoek, E and Brown, E T, Practical estimates of rock mass strength, International Journal of Rock Mechanics and Mining Sciences, 34(8), pp Louchnikov, V, 211. Ground Control Management Plan (Minerals and Metals Group). MMG, 21. Golden Grove Fact Sheet [online]. Available from < [Accessed: 5 October, 211]. Nickson, S D, Cable support guidelines for underground hard rock mine operations, Masters thesis (published), University of British Columbia, Vancouver.