Mining Factors Affecting Hanging Wall Dilution

Mining and rock mass factors influencing hangingwall dilution K. Forster & D. Milne University of Saskatchewan, Saskatoon, Canada

A. Pop Cameco Corporation Rabbit Lake Mine Site, Collins Bay, Canada

ABSTRACT: Many factors simultaneously influence stability and dilution in open stopes. This makes it difficult to assess the effect of individual changes to the rock mass condition or mining activity. Hangingwall stability is being studied at the Rabbit Lake Mine in Northern Saskatchewan. The influence of the rock mass condition, overall stope geometry and mining factors such as cable support are considered in this paper. Detailed analysis of 3-D stope survey data is used to accurately reflect the degree of instability for a given opening geometry, stress and rock mass condition. This stope survey data is coupled with detailed geology data to assess the influence of changing rock mass conditions away from the opening surface. This paper outlines an approach for gaining a better understanding of the factors influencing dilution based on detailed assessments of individual stopes, rather than a statistical assessment of many factors from a large data base of case histories.

1

INTRODUCTION

A key aspect of longhole open stope design is the prediction of the amount of overbreak or dilution that can be expected. Dilution is influenced by many factors controlled by both the mining and ground conditions. Mining controlled factors include, among other things, ground support, stope geometry, blasting method, stope sequencing, undercutting and time. In order to predict the dilution, some assumptions and generalities are required to estimate the rock properties. Underground openings in hard rock open stope mining are often designed using a well established procedure. Piteau (1973) states that the process begins with separating the rock mass into structural domains that are predicted to behave in a similar fashion or have similar properties. In general, these structural domains are relatively large, and they are based on lithological contacts or major zones of jointing or shearing. This generalized method is limited in that it was developed for assessing the stability of the walls in large open pits, and therefore was developed to analyze slope failures that were often hundreds of metres in extent. It was necessary to obtain average rock mass properties for large structural domains. In underground mining, the areas of concern are often only in the order of tens of metres, and therefore averaging rock mass properties over large areas may be too simplistic and lead to inaccurate prediction of stope behavior. In order to attempt to estimate the influence of both changing ground conditions and mining practice

on dilution, detailed records and stope reconciliation work must be collected and interpreted. Quantifying and interpreting these factors is often not straightforward.

2

The Cameco Corporation Eagle Point mine is located approximately 800 km north-east of Saskatoon, Saskatchewan. The mine produces an average of 600 tonnes of uranium ore per day using longhole, open-stope mining methods. 2.1 Geology Dishaw (2005) states that the uranium mineralization at the Eagle Point mine is structurally controlled, and typically occurs within the hangingwall of the Collins Bay reverse fault within a basal unit of metasediment. The host rock consists of both metamorphic rock and intrusives, in particular biotite-quartz-feldspar gneiss, quartz-feldspar gneiss, quartzite, calc-silicates and pegmatite. The mineralization zone is typically less than 1 m to 15 m thick, and consists of tabular veins and lenses dipping from 50 to 85 degrees. Due to the ore association with faulting, the ore is often contorted, with a highly variable orientation. Dishaw (2005) states that in addition to the uraninite ore, many other types of alteration assemblages are associated with the ore zone. Bleaching,

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BACKGROUND

clay or ‘argillization’, pyritization, hematization, carbonate ± quartz veins, quartz veins, sericitization and chloritization may be present in the ore zone in various different combinations. The alteration is typically confined to within 5 metres in the hangingwall of the ore zone, but can extend up to 20 metres into the hangingwall of high-grade mineralization, localized around multiple fault intersections. 2.2 Mining method As mentioned, longhole open-stope mining methods are used at the Eagle Point mine. Both narrow vein stopes and massive stopes have been used to extract the ore, depending on the width and spacing of the ore veins. The overcuts and undercuts for the narrow vein stopes are driven on geological control following single, or closely spaced ore veins. The stopes are then mined longitudinally with average strike lengths of 30 metres, and are generally less than 10 metres wide. The overcuts and undercuts for the massive stopes are driven on survey control through a series of ore veins that are too numerous and closely spaced to mine individually. The massive stopes are then mined either longitudinally or transversely, with an average strike length of 20 metres and widths generally less than 20 metres. 3 3.1

DILUTION PREDICTION Background

A widely accepted method of estimating stope design was developed by Potvin (1988). The dilution graph has been modified from this method by Clark (1998) and others, including Capes (2005) to accommodate stopes in weak rock. The stability graph is based upon two input parameters; the hydraulic radius, HR, and the modified stability number, N . The hydraulic radius is defined as the ratio of the area to the perimeter of a surface. The modified stability number, N is given as:

where Q = rock tunneling quality index; A = stress reduction factor; B = joint orientation factor; and C = gravity adjustment factor. Both the hydraulic radius and the modified stability number are influenced by several factors that need to be considered. 3.2

Difficulties in prediction method and analysis

The prediction of stope hangingwall behavior and analysis of actual stope dilution are very complex issues. There are many factors that influence the input data

Figure 1. Typical stope geometry showing undulating drifts and hangingwall design.

and output results. Uranium mining has the added constraint of minimizing radiation exposure to workers. The stopes are developed along the ore vein, with very little to no access to the hangingwall rock. Once a development round has been taken, it is shotcreted as soon as possible to limit gamma radiation exposure to the workers. This provides a limited time frame for personnel to inspect the rock mass and classify it appropriately. A method of supplementing the ore drift data collected by the mine geologists is required to refine the classification process for the purposes of design. 3.3 Hydraulic radius calculation Although the modified stability graph is designed for stope geometries, in reality stopes rarely fit the ideal model. A typical stope design is shown in Figure 1. The overcut and undercut drifts for the stope are not perfectly straight, and there is a fold in the hangingwall extent that is problematic in terms of stability. The most influential factor in the calculation of the hydraulic radius is how to account for support installed in the drifts. Cable bolt support is typically installed in most overcut and undercut drifts as shown in Figure 2. Cable support design methods exist which assess the effectiveness of an even coverage of cables (Potvin et al. 1989). This approach cannot be applied for hangingwall design in cases where support is only installed from sub-levels or an overcut and undercut drift (Potvin & Milne 1992). Methods have been developed for sub level support design for stope hangingwalls with multiple sub-levels (Nickson 1992). There are no design techniques for support design when support is only installed from the stope overcut and undercut. Three options to account for this support in the hydraulic radius calculation are shown in Figure 3.

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Figure 2. Typical cross section of a stope showing the drift outline, planned blast outline, cable bolts, results of the cavity monitoring survey (CMS), and geological structures.

Figure 3. Three potential measurements to calculate the hydraulic radius of the stope.

The first option is to assume the cable bolting will not affect the interpretation of the opening geometry. The support installed is effectively ignored and the drift openings and hangingwall are treated as one continuous surface. The second option is to assume the support installed provides some stability, but not as much as solid, undisturbed rock. This method assumes a hangingwall height measured from the middle of the overcut to the middle of the undercut. The third option is to assume the cable bolting provides sufficient support to treat the drift as solid rock. The hangingwall height is then assumed to be from the bottom of the overcut to the top of the undercut. This approach is taken by Hoek et al. (1995) where it is suggested that cable support from stope sub-levels will create a stable ‘buttress’ at each drift.

Figure 4. Cross section depicting the two angles that may potentially be used to calculate the gravity reduction factor, C.

3.4 Modified Stability Number (N’) calculation There are many separate factors which contribute to the calculation of the Modified Stability Number. The first term, Q , includes rock quality designation, number of joint sets, joint roughness, joint alteration, and joint water condition. As previously mentioned, there is limited access and limited time to gather geotechnical information for a stope. Data from diamond drill core provides several of the Q factors, while others, such as large scale roughness and joint water condition may be gathered or inferred from geological mapping and observations. The joint orientation factor, B, may change as the dip of the hangingwall changes. The B value may be averaged for the entire mining lens, or it may be assessed for an individually stope. The gravity reduction factor, C, is another difficult parameter to quantify. As shown in Figure 4, the angle used to calculate C depends on whether the stope design angle is used, or the overall stope angle as measured from the top of the overcut to the bottom of the undercut. For stopes that have a larger lift height, the difference in the angles is negligible. But for stopes with shallow-dipping hangingwalls, or shorter lift heights, the difference can be significant. 3.5 Other contributing factors There are other factors that influence the amount of stope dilution. It has been shown that the amount of time a stope is left open can significantly influence the amount of dilution that is produced (Wang et al. 2003). The Cavity Monitoring Survey (CMS) has been used to quantify the amount of stope dilution (Mah 1997). One drawback to relying on the CMS method for measuring dilution is the presence of shadows, which makes it difficult to accurately assess the amount of hangingwall dilution.

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Table 1. Correlation between the R1/R3 and A1/A7 geology system and Q’ classification systems (Sutton 1998).

Table 2.

Hangingwall condition.

Stope 200-045 data

Stability graph data

HW stress condition Orientation of structure

Hangingwall is relaxed A = 1.0 Major jointing makes an angle of about 10◦ to the hangingwall B = 0.2 Hangingwall dips at an average of 47◦ C = 8 − 6 cos θ = 3.9 Stope strike length = 32 m

Rock stregnth Alteration

R1

R2

R3

A1 A3 A5 A7

N/A N/A 1.2 0.4

N/A 3.8 2.5 2

22 11 8.3 3.8

There is no direct method to adjust the stability number N’ to account for discrete features such as faults, shears or weak zones with a strong degree of alteration (Potvin & Milne 1992). Work has been done to account for the influence of discrete features such as faults by quantifying the influence of these features on the induced stresses around openings (Suorineni et al. 1999). There has been limited work validating this approach through case histories. 3.6

4.1

Stope up-dip length

Hydraulic Radius

Current rock mechanics approach

Rock mechanics data has been gathered predominantly by the geology staff at the mine. The primary goal of mapping by the geology staff has, of course, been focused on following the ore for stope development. In 1997 a study was conducted by Sutton (1998) to link stope stability to the rock strength / alteration assessment developed by the geology staff. Available exposures were assessed into the alteration and strength categories and were also mapped for rock mechanics classification purposes. The mine currently uses the classification system based on the alteration and strength of the rock. The R series consists of three separate categories of rock strength, with R1 being very weak rock, and R3 being relatively strong rock. The A series consists of four separate categories for alteration, with A1 being fresh rock, to A7 being strongly altered. Sutton’s correlation between the Q’ rock mechanics classification system and the geology assessment of R1 to R3 and A1 to A7 rock types is applied to stope stability and is summarized in Table 1. The hangingwall hydraulic radius for supported overcuts and undercuts is currently being assessed by taking the up-dip extent of the stopes to the centre of the supported sub levels. The dip of the stope hangingwall is being taken as angle at an average of 47◦ . 4

Hangingwall orientation Hangingwall geometry

SAMPLE CASE STUDY Dilution prediction

The dilution graph has been used for an initial interpretation of the hangingwall dilution (Clark & Pakalnis

1997). The stope properties for Stability Number N’ and Hydraulic Radius calculation are summarized in Table 2 (Potvin 1988). As an initial assessment, the drift mapping data in Figure 5 indicates an A5/R2 alteration/strength rating, which suggests a Q’ value of 2.5 which allows for the following Stability Number calculation:

As shown in Figure 6, the modified stability graph predicted more dilution than was measured using the cavity monitoring survey. Although the influence of shadows caused by bends in the hangingwall for the rings closer to the slot may have shown less dilution than actually occurred, it is believed that the average dilution for the stope was still not more than 0.5 metres. This indicates that the method and parameters that were used to predict the dilution were pessimistic in this particular case. Upon inclusion of detailed core logging data as shown in Figure 7, it appears that the stope sloughed to a geological contact that was not apparent in the geological drift mapping. The cross section shows

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Top of overcut to bottom of undercut Length = 21 m Centre of overcut to centre of undercut Length = 18 m Designed hangingwall Length = 13.5 m Top of overcut to bottom of undercut (32 m)(21 m) = 6.3 m HR = 2(32 m + 21 m) Centre of overcut to centre of undercut (32 m)(18 m) HR = = 5.8 m 2(32 m + 18 m) Designed hangingwall (32 m)(13.5 m) HR = = 4.7 m 2(32 m + 13.5 m)

Figure 5. Plan view showing stope 200-045 layout and geological mapping of the overcut and undercut drifts.

recorded in terms of percentage of individual alteration products as well as a general, overall alteration. The alteration may be found on joint surfaces as well as the rock fabric may be altered throughout. The scale used to denote alteration in core logging is based on 0 for fresh rock, 1 for weakly altered rock, 2 for moderately altered rock, and 3 for strongly altered rock. It is suspected that the pegmatoid rock is less stable at similar alterations and rock quality designation than the gneiss. Work is ongoing to check this assumption. Also, it is possible that the cable bolting had more of an effect on the stability of the stope than was accounted for in the design process. The assumption that a cable supported overcut / undercut provides as much support as a rock abutment seems overly optimistic. Ignoring the cable support does not seem realistic. Continued work in this area, with emphasis on the hangingwall profile from the CMS in the vicinity of the cabled drifts, will be investigated with additional case histories. Figure 6. Modified stability graph showing the three calculated hydraulic radii and the predicted dilution.

the geological back mapping as well as one of the exploration drillholes. The lithology, rock quality designation, and degree of alteration are shown on the section. During the core logging process, the alteration is

5

As discussed, there are many individual factors that can influence the hangingwall stability. The complex interaction of these factors makes it difficult to predict the hangingwall behavior. In addition to the factors that

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CONCLUSIONS

Figure 7. Stope cross-section showing where the hangingwall sloughed to a geological contact.

have been previously identified, there is the issue of complex geology to be considered. More research and analysis are required to better quantify all of these factors. Further study is currently being conducted to investigate the effects of complex geology and develop a method of dilution prediction from exploration core data. REFERENCES Capes, G., Milne, D. & Grant, D. 2005. Stope hangingwall design approaches at the Xstrata Zinc, George Fisher Mine, North Queensland, Australia. In U.S. Rock Mechanics Symposium, Fairbanks. Clark, L. & Pakalnis, R. 1997. An empirical design approach for estimating unplanned dilution from open stope hangingwalls and footwalls. In CIM AGM, Calgary. Dishaw, G.R. 2005. Rabbit Lake Operation: Canada’s longest operating uranium mine – 30 years and still glowing. CIM-AGM, Toronto. Hoek, E., Kaiser, P. & Bawden, W. 1995. Support of Underground Excavations in Hard Rock. Rotterdam: Balkema, pp. 215. Mah, S. 1997. Quantification and Prediction of Wall Slough in Open Stope Mining Methods. MASc Thesis, University of British Columbia, 290 pp.

Nickson, S. 1992. Cable support guidelines for underground hard rock mine operations. M.A.Sc. thesis, University of British Columbia, 223 pp. Piteau, D.R. 1973. Characterising and extrapolating rock joint properties in engineering practice. Rock Mechanics 2: 5–31. Potvin,Y. 1988. Empirical open stope design in Canada. Ph.D. thesis, University of British Columbia, 350 pp. Potvin,Y., Hudyma, M. & Miller, H. 1988. Design guidelines for open stope support. CIM Bulletin, June. Potvin, Y. & Milne D. 1992. Empirical cable bolt support design. In Proceedings Rock Support in Mining and Underground Construction. Rotterdam: Balkema, pp. 269–275. Suorineni, F., Tannant, D. & Kaiser, P. 1999. Fault factor for the stability graph method of open stope design. Trans. Instn Min. Metall. (Sect A: Min. Industry) 108. Sutton, D.A. 1998. Use of the Modified Stability Graph to predict stope instability and dilution at Rabbit Lake Mine, Saskatchewan. University of Saskatchewan Design Project, Canada. Wang, J., Milne, D., Yao, M., Allen, G. & Capes, G. 2003. Open stope exposure time and stope dilution. In CIM AGM, Montreal.

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