1 INTRODUCTION

Developing a mining project depends on technical, economical, political and societal feasibility. This is reflected in the staged approach commonly adopted for developing mining projects. These stages typical-ly start with preliminary studies followed by feasi-bility designs and the definition of a mining plan. Even at preliminary stages of mine design completed before significant geotechnical investigations are done, stakeholders require estimates of capital and operational costs for decisions to be made on the vi-ability of the operation. This implies that prelimi-nary mine designs are needed once an ore body has been identified and an initial characterization of its mineralogy and extent is achieved. In most situa-tions, this means a limited budget for geomechanical exploration. As a consequence, the geomechanical rock mass classification, characterization and exca-vation designs at these preliminary stages often rely mostly on rock core obtained for mineral exploration purposes. The density and coverage of mineral ex-ploration boreholes also make them a valuable source of geomechanical information at advanced stages of design.

However, using mineral exploration core for geo-mechanical characterization of a rock mass has sev-eral important limitations. Original core logs of the-se boreholes focus on geological descriptions and omit important geomechanical information about the intact rock core such as joints, shear zones and other structures. Re-logging these cores and sampling for geo-mechanical laboratory testing become neces-sary. Assessment of the results should consider ef-fects from disturbance and alteration of the materials due to core handling and from changes during stor-

age. In this regard, the information obtained from exploration boreholes and the excavation designs based on this information will have an associated level of uncertainty.

The designer should understand how uncertainties and biases related to geomechanical characteriza-tions based on re-logged cored influence the design, and their importance relative to other sources of un-certainty.

Morgenstern (1995) divided geotechnical uncer-tainty into three categories (Figure 1): a) Model Un-certainty, related to our limitations in the theory and models used for prediction of ground behaviour; b) Parameter Uncertainty, related to influences on the inputs used for analysis; and c) Human Error, related to people’s mistakes (El-Ramly 2001).

The uncertainties examined in this paper are re-lated to rock mass characterization based on re-logging stored boreholes, and correspond to the Pa-rameter Uncertainty type.

Figure 1 shows how Parameter Uncertainty is fur-ther subdivided (After Baecher 1987 and El-Ramly 2001). The subdivision corresponds to sources of Parameter Uncertainty. Detailled discussion of these can be found in Baecher (1987). Data Scatter corre-sponds to the variability of measures around a cen-tral tendency and Systematic Error corresponds to deviations of the measured data trend from its real trend. Data Scatter is divided into Real Spatial Vari-ability of the data and variability induced by limita-tions on sampling and testing (Random Testing Er-ror). Systematic Error is divided into Statistical Error, associated with how representative the sam-ples are in statistical terms; and Bias in Measure-ments, related to overestimations or underestima-

Empirical Design of Underground Excavation Support: Assessing Reliability Associated with Rock Mass Classifications using Core Drilled for Mineral Exploration

Renato Macciotta Department of Civil and Environmental Engineering, University of Alberta, Canada

Graham Parkinson Klohn Crippen Berger Ltd., Vancouver, Canada

Eduardo Garcia & Orlando Bravo Klohn Crippen Berger Ltd., Lima, Peru

ABSTRACT: Design of underground mining operations, in its preliminary stages, requires obtaining geomechanical data from re-logged rock core initially drilled for mineral exploration. The degree of alteration of this core due to handling and storage time increases the uncertainty in the geotechnical parameters assessed and in the estimated required excavation support. This paper presents a simple approach to quantitatively assess this uncertainty, illustrated with a case study in the Peruvian Cordillera. The Q-System was adopted for a preliminary assessment and to estimate the required excavation support.

tions of parameters due to testing apparatus, bounda-ry conditions, disturbance, correlations, etc.

The changes in the geomechanical characteristics of rock core stored for several years can systemati-cally or randomly deviate the trend of measured ge-omechanical parameters.

This paper examines how identified sources of uncertainty could possibly correspond to systematic Bias in Measurements (Figure 1). One example is how ore exploration focuses on sampling the miner-alized zones and also disturbs (with removal of core, splitting of core) the geomechanical sampling of the-se sections within the exploration boreholes. This makes characterization of these zones to be chal-lenging and increases the measurement bias. The missing sampling of differences in the rock mass quality within the ore zone would tend to distort the overall distribution of rock mass quality within the rock unit. This bias needs to be understood, quantita-tively or qualitatively, to increase the reliability of underground excavation designs using such infor-mation.

This paper presents a case study where both five year old and freshly drilled ore exploration bore-holes were used for estimation of underground exca-vation support of access adits, ramps and service caverns for a mine operation in the Peruvian Cordil-lera. The two drillhole data sets allowed comparision of the effects of core storage on geomechanical par-amaters. The existence of an exploration adit in the area drilled provided further control.

The analysis was done at a preliminary stage and was based on the Q-System for rock mass classifica-tion and support design (Barton et al. 1974, Grim-stad and Barton 1993, Barton 2002). The Q-System was selected because of the vast experience and lit-erature on empirical excavation support using this system. Furthermore, it considers rock mass dis-aggregation, weathering and joint surface character-istics, among others. It was noted that some of these are likely to change during the time the core was ex-posed to the environment while stored. However in this case fresh core was also available from infill holes which presented an opportunity to systemati-cally compare the effects of changes with the fresh core used as a control.

This case study compares parameters obtained for estimation of the Q value: Rock Quality Designation (RQD), Joint Roughness Number (Jr), and Joint Al-teration Number (Ja). The comparison was done be-tween exploration boreholes drilled in 2007 and 2012. A qualitative and quantitative assessment of the variability of these parameters is presented. The assessments were adopted as a means to assess the Parameter Uncertainty associated with the use of 5-year-old stored rock core for preliminary estimations of underground excavation support.

Figure 1. The three main sources of uncertainty in geotechnical design (After Morgenstern 1995) focused on Parameter Uncer-tainty (After Baecher 1987, and El-Ramly 2001).

2 STUDY AREA

2.1 Project location The mining exploration project is located in Amazo-nas, Peru; about a 5-hour drive (300 km) from the city of Tarapoto. Figure 2a shows the location of the project.

The area is characterized by its rugged topogra-phy, steep mountainous terrain, and dense vegetation (Figure 2b). The vegetation is typical of the eastern section of the Andes Cordillera in Peru, where pre-cipitation exceeds 1000 mm/year and temperatures exceed 30 degrees Celsius.

Figure 2. Location of the study area (a) and photograph illus-trating the terrain and vegetation in the area (b).

Parameter Uncertainty

Data Scatter Systematic Error

Real Spatial

Variability

Random Testing

Error

Statistical

Error

Bias in

Measurements

Data Scatter Systematic Error

Y

X

Y

X

Measure

d trend

Real trend

Sources of Uncertainty

Model Uncertainty Human Error

Pacific

Ocean

Peru

Study

area

a)

b)

2.2 Geology and Weathering Within the projected underground workings, the rock units present are dominated by limestones and dolomites of the Chambara formation. The Chamba-ra formation can be subdivided in three different groups. These are named Chambara 1, 2 and 3; in order from earlier to later time of sediment deposi-tion, where Chambara 3 overlies Chambara 2, which overlies Chambara 1. The ramps and service caverns projected lie within Chambara 2 and 3. Chambará 2 consists of dolomites and limestones. These units host the mineralized body and have an average thickness of about 200 m. Chambará 3 consists mainly of moderately to high bituminous limestones, and cherts can be found in some areas. Chambará 3 has an average thickness of 250 m. Fault or joint in-fill is typically composed of bitumen or clays. Moreover, these units are characterized by the pres-ence of karstic caverns, which tend to follow the ma-jor faults and define the hydrogeologic regime in the area. The sedimentary origin of these formations and the orogenic episodes that caused the folding and subsequent formation of an anticline in the study ar-ea; provide a context of persistent discontinuities parallel to bedding, non-persistent jointing near per-pendicular to bedding, and faulting.

The rock core and the joint infill materials would have undergone different degrees of weathering in-situ and while stored, depending on the storage time and the exposure to the environment. Weathering is the process of alteration and breakdown of rock and soil materials at and near the Earth's surface by chemical decomposition and physical disintegration (Anon 1995). The type of weathering process and the materials product of it highly depends on climate and lithology. Useful reviews of weathering pro-cesses are presented by Carrol (1970), Ollier (1986), Yatsu (1988), Selby (1993) and Price (1995).

Due to the varied processes leading to weathering and rock lithology, no universally valid generaliza-tion can be done regarding the weathering products of different rocks. However, some common charac-teristics can be observed. In limestones and dolo-mites, when the presence of soluble minerals is the dominant factor, karstic weathering may occur (Price 1995). These type of weathering has been en-countered through the study area, as will be shown later in this paper. In these types of rock, minor changes in solubility or discontinuity distribution and characteristics, can lead to differences in the size and distribution of karst cavities. Differences in the content of calcite and dolomite within Chambara 2 and 3 can potentially modify the degree of weather-ing each undergo when stored.

Influence of other weathering processes can also lead to the formation of other clay minerals within discontinuity surfaces, such as montmorillonites. Preferential water paths defined by jointing and

faulting would have influenced the formation of karstic caverns, but also allowed for external agents to get in contact with the rock mass. Several karstic openings and joint walls within the study area had clay coatings. Also, some joints presented clay infill up to 10 to 15 cm thick. The clay infill in the rock core was altered due to desiccation during storage with resulting loss of infill due to fragmentation of the desiccated infill material.

The characteristic infill of bitumen associated with the deposit also underwent changes during core storage. Bitumen that was found infilling some of the discontinuities was also desiccated in the rock core and undergoes loss of volatiles which altered it’s frictional characteristics. Bitumen observed in-situ had lower friction angle and was much softer than in desiccated cores.

3 ROCK MASS QUALITY OBSERVATIONS

Rock mass quality observations were based on rock outcrops in the area, the 2007 and 2012 drill core samples, and observations in an exploration tunnel (3.0 m by 2.5 m opening section and 5 m to 6 m wide drill chambers). Of particular interest are the differences in the rock mass characteristics observed between the exploration tunnel and the drill core. This can provide a qualitative assessment of the rock mass quality changes induced by core handling and storage.

Figure 3a and Figure 3b show two rock outcrops in the study area. Common to other sedimentary formations, persistent bedding is present. This bed-ding has lead to persistent discontinuities with varia-ble spacing between them, as shown in Figure 3b. Outcrops typically showed a stratified rock mass, with competent rock matrix between these persistent sub-horizontal discontinuities.

Observations during the tunnel site visit indicate that tunnelling conditions in the areas of the Cham-bara 2 and 3 are generally good. No support was re-quired in the excavation of the tunnel sections. Ad-vance of the tunnel was achieved using 3-foot to 6-foot rounds. Drill chambers were supported success-fully by 6-foot split sets at 1.5 m to 2.0 m spacing, requiring between 3 and 5 split sets per chamber. Figure 3c shows a typical section of the tunnel. The-se observations suggest a cohesive, competent rock mass that is stable under these excavation dimen-sions with little or no support.

Figure 3d shows a zone of concentrated defor-mation, leading to a zone of shearing where the rock mass decreases in quality. These observations were supported by the drill core observations. Figure 4a shows a section of core with high RQD value and a shear zone approximately 1.5 m thick and highly weathered. The rock mass surrounding the shear zone does not show significant alteration, but does

have an elevated presence of joints. Zones of lower quality rock mass are also present. An example is shown in Figure 4b, where a shear zone about 1.2 m thick is surrounded by increasingly jointed and al-tered rock.

Discussions with site personnel indicate that con-trol of the face was not a problem during tunnelling. Several notable faults were intersected that required timbering and shoring to prevent ravelling of loose materials (Figure 3e). Indicators of motion such as slickenside striations are present in the core and were observed in the exploration tunnel. These are often in multiple phases denoted by multiple move-ment planes within discontinuities. No signs of over-stressing (rock bursts, rock fall, spalling, defor-mation) were observed within the tunnel sections. Simple visual inspection down large diameter un-derground boreholes through open collars did not show any signs of overstressing (i.e. breakouts with-in borehole walls) and the core did not exhibit disc-ing. This suggests that the rock mass withstands ex-cavation induced stresses under the in-situ state of stress.

These comparisons between the observations in the exploration tunnel and the stored core suggested that the rock mass quality did not vary significantly due to rock core storage. Some rock core disaggre-gation can be expected from handling and weather-ing, as well as losses in moisture content. Identifica-tion and marking of the natural insitu joints versus mechanical breaks caused by handling or core ex-traction was a valuable component of the geome-chanical core re-logging program.

Figure 3. Observation inside the exploration tunnel. Rock out-crops in the study area (a and b), typical section of the explora-tion tunnel (c), zone of concentrated displacement (d), fault zone (e), and karst cavern (f).

Figure 4. Observations from drill core. a) Shear zone surround-ed by new (2012), high RQD core, b) 5-year-old (2007) core with more disaggregated and altered mass surrounding a shear zone, c) signs of dissolution, d) bitumen and e) clay infilling discontinuities.

As previously discussed, loss of moisture during

storage would have a more significant effect on the materials infilling the discontinuities. The desicca-tion process altered these infill materials into stiff and brittle ones, in contrast to their soft and ductile behaviour observed in the exploration tunnel. How-ever, it is judged that stress relief induced by tunnel-ling and concentration of groundwater flows toward the tunnel could have lead to joint opening and al-lowed the infill material to take more moisture, therefore softening when compared to natural condi-tions. The infill materials are then expected to pre-sent slightly stronger characteristics than those ob-served in the exploration tunnel, however still relatively soft and ductile. It is important to note however that their characteristics as observed in the exploration tunnel are those as to be expected around the mining openings. Fresh infill material from the exploration tunnel was excavated, sampled and tested to compensate and compare against the weathering of the infill in the core used for logging.

Karst within the tunnel was associated with the presence of opened discontinuities. Minor karst was usually associated with minor fault zones, usually a few centimetres thick (20 to 40 cm). Large caverns (Figure 3f) were associated with thicker fault zones (up to 2 metres) with openings of up to 0.5 m, and infill materials that ranged from blocks to coble siz-es, and within a matrix of high plastic clayey materi-al and bitumen. Signs of dissolution were also obvi-ous in the drill core (Figure 4c), as well as the presence of bitumen and clay infilling some of the discontinuities (Figure 4d and Figure 4e).

Although the tunnel is generally dry, water is en-tering the tunnel through a variety of karst openings above the tunnel and flowing out of the tunnel to levels below. Three main flows were observed, total-ling less than 3 L/s. Karst systems communicate with the surface and inflows increase during periods of heavy rainfall. These observations were used to assess how the presence of water influences the quality and strength of the rock mass using the Q-System, as presented later in this paper.

4 MEASURED RQD, JR AND JA

The preliminary geomechanical characterization of Chambara 2 and 3 units had two objectives: 1) pro-vide the basis for estimating the support required for permanent structures (adits, ramps and service cav-erns), and 2) provide preliminary parameters for an initial assessment of the extraction volumes and di-mensioning of the mining operations. Hereafter the paper will focus on the first objective.

The information available for the preliminary characterization consisted on the observations from the site visit, and re-logged drill core originally ob-tained for mineral exploration. The oldest drill core selected had been stored initially outdoors in boxes and later indoors in covered racks for over 5 years (drilled in 2007), and the newest core had been stored under cover for only a few months (drilled in 2012). Given the high temperatures and humidity in the area, it was a concern that the oldest core would have been significantly altered, consequently in-creasing the uncertainty in the geomechnical charac-terization. A comparison of rock mass quality pa-rameters between the oldest and newest re-logged core is presented in this paper to quantify and mini-mize this uncertainty.

An ideal scenario for comparing geomechanical characterizations is to consider the values obtained at the same locations and depths. However this was not achievable however infill drillholes were availa-ble in similar areas. Drilling and Re-logging bore-holes at the same locations would have proven cost-prohibitive. Using all of the measured borehole val-ues for comparison gave good statistical coverage and was considered appropriate given the proximity of boreholes and that visual inspection of the core suggested the conditions of Chambara 2 and Cham-bara 3 do not vary significantly across the study ar-ea.

4.1 Measured Rock Quality Designator (RQD) RQD (Deere et al. 1967) is defined as the percentage of intact core pieces longer than 10 cm. It was cho-sen for comparison because it can be taken as an in-direct indicator of the degree of rock mass disaggre-gation. Figure 5 shows the histograms of measured

RQD for Chambara 2 (Figure 5a) and Chambara 3 (Figure 5b) based on more than 500 core samples.

These histograms compare the measurements done with the core drilled in 2007 and 2012. Figure 5 compares the measured RQD of the entire drill core. It is expected that lower values would corre-spond to shear zones and karst areas, with limited thickness, zones that require timbering and shoring to prevent ravelling of loose materials.

Visually, the comparison shows good correlation between the ranges of measured RQD between the old and new drill core. It was then decided to adopt the Correlation Coefficient between these data to quantify how good this correlation is.

The Correlation Coefficient is a measure of the linear correlation between two variables. It can take a value between +1 and −1, inclusive. A value of 1 is perfect positive correlation, 0 is no correlation, and −1 is perfect negative correlation. Its mathematical definition can be found in Sharma (2005) and most introductory textbooks on statistical data analysis. The Correlation Coefficients between the old and new drill core values of RQD were 0.90 and 0.76 for Chambara 2 and Chambara 3, respectively. These were considered to represent good correlations be-tween the data. This suggests that in limestones and dolomites there is little change in RQD as a result of storage of the core.

Figure 5. Histogram of measured RQD for Chambara 2 (a) and Chambara 3 (b). Light grey corresponds to rock core obtained in 2007 and dark grey to core obtained in 2012.

4.2 Measured Joint Roughness and Joint Alteration (Jr and Ja) The measured Q parameters Jr and Ja are of interest in assessing effects of storage because they describe the condition of the discontinuity walls, contacts and infilling materials, which could be altered by weath-ering, dessication and loss of infill due to vibration caused by core handling during storage. The defini-tion and discussion of these parameters can be found in Barton et al. (1974), Grimstad and Barton (1993) and Barton (2002). Figure 6 shows histograms of measured Jr for Chambara 2 (Figure 6a) and Cham-bara 3 (Figure 6b). These histograms compare measurements done with the core drilled in 2007 and 2012. The corresponding histograms for the meas-ured Ja are shown in Figure 7a (Chambara 2) and Figure 7b (Chambara 3).

Figure 6a and 7a suggest good correlations for Jr and Ja in the Chambara 2. This is supported by Cor-relation Coefficients of 0.99 and 0.95 for Jr and Ja, respectively. Jr for Chambara 3 shows a good corre-lation (Correlation Coefficient of 0.85 – Figure 6b). However, some measures of the old core seem to be shifted towards lower values (from Jr = 3 towards 1). This could correspond to rough joint surfaces (Jr = 3) being altered towards a smoother surface by weathering and handling, both of which could cause erosion of asperities resulting in increased infill or crushed materials (Jr = 1). This implies that Jr meas-ured with old core could tend to underestimate the in-situ rock mass quality by a factor of 0.33. How-ever, this is not considered significant given that the values of Q can range within 6 orders of magnitude.

Figure 6. Histogram of measured Jr for Chambara 2 (a) and Chambara 3 (b). Light grey corresponds to rock core obtained in 2007 and dark grey to core obtained in 2012.

Figure 7. Histogram of measured Ja for Chambara 2 (a) and Chambara 3 (b). Light grey corresponds to rock core obtained in 2007 and dark grey to core obtained in 2012.

Ja for Chambara 3 shows a lower correlation be-

tween older and fresh core (Correlation Coefficient of 0.65 – Figure 7b). The measures of the old core also seem to be shifted towards lower values (from Ja = 4 towards 2). This could correspond to discon-tinuities with clay mineral coatings, sand particles and disintegrating rock (Ja = 4) experiencing losses of the infilling material while handling, or hardening of infill due to moisture loss; with an apparent im-provement in joint wall conditions (Ja = 2). This im-plies that Ja measured with old core could tend to overestimate the rock mas quality by a factor of 2. Again, this is not considered significant given the values of Q can range within 6 orders of magnitude.

5 ESTIMATED Q AND EFFECTS OF WEATHERING OF CORE ON EXPECTED RANGE OF REQUIRED EXCAVATION SUPPORT

Design of excavations on stratified rock masses need to consider the response of the rock mass to excava-tion. The persistent sub-horizontal bedding in the ar-ea suggests that the excavation roof will behave as a roof beam, and roof design should consider this be-haviour (Brady and Brown 2004). Given the signifi-cant available experience using the Q-System for underground support excavation, Q was adopted in this paper for a preliminary estimation of the re-quired support of the permanent underground exca-vations (adits, ramps, and service caverns).

5.1 Estimated Q Q is calculated with the following expression (Bar-ton et al. 1974):

𝑄 =𝑅𝑄𝐷𝐽!

𝑥𝐽!𝐽!𝑥𝐽!𝑆𝑅𝐹

Where RQD, Jr and Ja were previously defined,

Jn is the Joint Set Number, Jw is the Joint Water Reduction Factor, and SRF is the Stress Reduction Factor. Jn and SRF are not expected to be highly sensitive to drill core being stored if RQD showed no sensitivity. SRF, however, might slightly vary in a similar fashion to the inferred variations in infill when assessing Ja and Jr within Chambara 3. Jw is a parameter that depends on the in-situ conditions at the time of drilling and was assumed between 0.3 and 0.7 (large to low water inflow, which varies sea-sonally). These parameters were adopted based on the observations within the exploration tunnel dis-cussed previously. Jw is a parameter that is difficult to assess when using historical exploration drill core without detailled drilling records and typically has to be assessed based on an understanding of the groundwater conditions. Also, it is worth to note that RQD / Jn does not change much over time for core in storage and that is typically the parameter with the largest numerical range.

Figure 8 shows the histograms of estimated Q for Chambara 2 (Figure 8a) and Chambara 3 (Figure 8b). These histograms compare the measurements done with the core drilled in 2007 and 2012. It is ex-pected that the lower values of Q correspond to shear zones and karst areas, limited in extent, and that require timbering and shoring to prevent ravel-ling of loose materials.

Figure 8 suggests a good correlation between the Q values estimated with the old and new drill core. The Correlation Coefficient between the Q values estimated with the old and new drill core is 0.95 for both Chambara 2 and Chambara 3. This suggests that the variations found for Jr and Ja in Chambara 3 between old and new core as a result of weathering are not exerting significant influence in the overall predicted rock mass quality estimations (Q values). This can be a consequence of both the small magni-tudes of the effects of weathering and that the mag-nitude of these variations or that variations in Jr and Ja tend to balance when calculating Q.

5.2 Expected Range of Required Excavation Support

The required excavation support was estimated following Grimstad and Barton (1993). It was con-sidered that most values of Q are at or above 1, and that lower values correspond to shear zones with

limited thickness (up to 2 m) that will require tim-bering and shoring.

The support estimation was done for adits and ramps with a maximum opening dimension of 3 m, and for service caverns with a maximum opening dimension of 8 m. The Excavation Support Ratio (ESR in Figure 9) was adopted as 1.6, which corre-sponds to the suggestion of Barton et al. (1974) for permanent mine openings.

Figure 9 shows the expected excavation support for the underground openings. According to the ap-proach adopted in this paper, adits and ramps with a maximum opening dimension of 3 m are expected to be stable when unsupported. This is consistent with the observations at the exploration tunnel. Most cav-erns with maximum opening dimension of 8 m with-in the Chambara 2 are also expected to be stable when unsupported, however at some locations with-in Chambara 2 and within most Chambara 3, these caverns would require systematic bolting (1.7 to 2 m spacing) and up to 50 mm of shotcrete.

6 DISCUSSION AND CONCLUSION

A proactive approach for managing uncertainty in such as quantitatively assessing influences of the use of weathered core on underground excavation design increases its reliability.

Figure 8. Histogram of estimated Q for Chambara 2 (a) and Chambara 3 (b). Light grey corresponds to rock core obtained in 2007 and dark grey to core obtained in 2012.

This, together with an observational approach

during operation of the mine, is considered a best practice towards a safe and efficient mining opera-tion. This paper has presented an approach to quan-tifing the uncertainty related to using stored drill core for geomechanical characterization of limestone and dolomite rock masses within Mississippi Style deposit environments. The case study estimated this uncertainty by comparing rock mass quality parame-ters (RQD, Jr and Ja) obtained from old and new drill core.

The approach presented in this paper was used for preliminary empirical estimations of underground support, but it can be expanded to advanced stages of design and the adoption of other design methods. Also, this approach can be extended to other pa-rameters such as Point Load Index and Unconfined Compressive Strengths. As an example, Point Load Index (PI50) measured in 2013 from rock core ob-tained in 2007 and 2012 are presented in Figure 10 and Figure 11 for Chambara 2 and 3, respectively. These comparisons follow the same approach as dis-cussed for RQD, Jr, Ja, and Q. A preliminary as-sessment of these distributions suggest the normal larger scatter of the data but with a similar central tendency in the estimated PI50. However, based on this data, no conclusive statements can be made at this moment and more data would be necessary to clearly identify any change in measured PI50 be-tween new and old core as a result of weathering.

Similarly, Figure 12 show Unconfined Compressive Strength (UCS) measured in Chambara 2 and 3 in rock core samples tested in 2013 obtained from boreholes drilled in 2007 and 2012. This figure sug-gests a large scatter in UCS for Chambara 3 when compared to Chambara 2, however only 4 samples were tested for Chambara 2 and this scatter is not considered representative of the unit. The UCS val-ues estimated for both units doesn’t show significant differences between tests in old and new core. This analysis is also considered preliminary and more testing is required to clearly identify any trend. As such, these figures are included for illustrative pur-poses only.

Figure 10. Point Load Index (PI50) measured in Chambara 2 in 2013 from rock core obtained in 2007 and 2012.

1 2 3 4 5 6 7 8 9 10PI@50D

0.1

0.2

0.3

0.4

Normalized Frequency

Chambara 2

2007 data 2012 data

Figure 9. Expected excavation support for adits and ramps with maximum opening dimension of 3 m and caverns with maximum opening dimension of 8 m. Expected support ranges are plotted on the estimated support categories based on the tunnelling quali-ty index Q (After Grimstad and Barton 1993).

Figure 11. Point Load Index (PI50) measured in Chambara 3 in 2013 from rock core obtained in 2007 and 2012.

Figure 12. Unconfined Compressive Strength (UCS) measured in Chambara 2 and 3 in rock core samples tested in 2013 ob-tained from boreholes drilled in 2007 and 2012.

In the case of the mainly limestone units ob-served, good correlation was found between the rock mass quality parameters derived from old core and new core, for the case study presented in this paper. This increases the confidence in the estimated values of Q, which in turn increases the confidence in the estimated excavation support. In this regard, the un-certainty in rock mass quality is mostly attributed to the natural variability of the rock mass, and the use of old core, in this case, is not a significant source of uncertainty. It should be noted, however, that un-measured bias exists related to missing and disturbed core within the ore zone. This corresponds to effects from core sampled for mineral exploration before geomechanical characterization is done. In this re-gard, the results presented in this paper are applica-ble only to the design of adits, ramps and service caverns; and were not directly extrapolated to the ore zones without engineering judgment. In this case, observations suggested that in general the ore zone has a lower rock mass quality. This should be considered in the preliminary dimensioning and support designs of mines when based on geome-chanical logging exploration core, and confirmed with geotechnical drilling in subsequent stages of design.

ACKNOWLEDGMENTS

The authors would like to acknowledge the team of specialists at Klohn Crippen Berger Ltd. involved with the logistics, field work and desktop studies as-sociated with the work presented here.

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0.4

0.5

0.6Normalized Frequency

Chambara 3

2007 data 2012 data

CH22007

CH22012

CH32007

CH32012

50

100

150

200UCS HMPaL