Keynote Lecture: Improving seismic risk management in hardrock mines

IMPROVING SEISMIC RISK MANAGEMENT IN HARDROCK MINES - 371 -

KEYNOTE LECTURE: IMPROVING SEISMIC RISK MANAGEMENT IN HARDROCK MINES

Y. Potvin and J. Wesseloo Australian Centre for Geomechanics, the University of Western Australia, Australia

The possibility of experiencing a seismic event resulting in fatalities has arguably become the most important financial risk in underground hardrock mines operating in developed countries. In the two most recent cases in Australia, the entire operation was shut down for a period well exceeding one year while the mining method had to be re-engineered in order to demonstrate to regulators that the seismic risks had been lowered to an acceptable level. The methods for mitigating seismic risk have developed significantly over the last couple of decades. The seismic risk is mitigated by reducing the seismic hazard through the implementation of sound geotechnical principals in mine design or through pre-conditioning practice, managing exposure of personnel and assets, and reducing the damage potential with dynamic resistant support. This paper reviews the risk mitigation techniques currently used in Australia, Canada and Sweden (A/C/S) with an emphasis on where and how the authors believe these techniques could be improved through future research and development.

INTRODUCTION

Mine seismicity and, more specifically, the possibility of experiencing a seismic event resulting in one or multiple fatalities has arguably become the most important financial risk in underground hardrock mines operating in developed countries such as Australia, Canada and others. This is in the context of the general definition of risk, which is the product of the hazard multiplied by the consequences.

Recent history in Australia has shown that the consequences of a fatality resulting from a seismic event are generally far more severe than any other type of fatalities. Common causes of underground fatal accidents other than seismicity, such as for example the ones related to mobile equipment or falling from heights, would generally result in the mine or a part of the mine shutting down for a period of days or a few weeks at the most. The last two fatal accidents related to seismic events in Australia have occurred, respectively, in 2006 at the Tasmania Mine (named the Beaconsfield Gold Mine at the time) and in 2000 at the Big Bell Mine. Both cases resulted in shutting down the entire operation for a period well exceeding one year. In both cases, the mining method had to be re-engineered in order to demonstrate to regulators that the seismic risks had been lowered to an acceptable level.

The financial consequence of a fatal accident resulting from a seismic event are likely to involve a long term shut down, and this cost alone will generally far exceed the total cost associated with any other type of fatality in underground mines.

In that context, it is not surprising that the management of seismic risk has become a priority in many of our deep underground mines. A number of measures can be adopted to manage seismicity. Amongst them, the use of seismic monitoring systems has become almost essential to understand the causes of seismicity and to manage the risk.

There is also often a need for up-grading the ground support practices to cater for dynamic loading.

In this paper, mitigation techniques such as seismic monitoring, dynamic support and others currently used in A/C/S and other developed countries for managing seismic risk will be discussed with an emphasis on where and how the authors believe these techniques could be improved through future research and development.

SEISMIC MONITORING

The following comments on the application of seismic monitoring in mines are included to provide a perspective on some of the ideas put forward in this keynote.

In countries such as A/C/S, the application of seismic monitoring techniques have many commonalities when compared to the approach historically used in South Africa, where many of the mine seismic monitoring equipment and analysis techniques have been developed. In particular, the same or similar seismic monitoring equipment is installed and many of the analysis techniques applied are the same. However, there are also some significant differences in the approaches to the problem. This is, at the origin, largely due to the differences in the geological settings and mining methods.

The South African deep level gold mines generally mine a thin, relatively shallow dipping reef in sedimentary deposits, and although three dimensional complexities occur as a result of reef displacement due to faults, generally the mining can be simplified to a two dimensional geometry.

Mining is not highly mechanised and the mine face advances with small blasts at relatively slow but constant rates. Seismicity in the deep level gold mines in South Africa started to become a significant problem in the 1960s and a lot of effort has been applied to the problem since then. Very valuable inroads into the problem have been

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made and when the first seismic system was installed in Australia in the mid 1990s at Mount Charlotte, Australia had the benefit of 30 years of research and development from South Africa. The mining environment is different however, and as a result, not all the knowledge gained in South Africa was directly transferable.

In A/C/S, the geology and ore deposits can be very complex and three dimensional. Mining advances rapidly with large blasts causing sudden changes in stope geometry and drastic stress redistributions. Complex sequences of extraction are often devised.

In the South African environment, due to the very large but regular two dimensional mining area that needs to be covered, the density of sensors used in their typical seismic monitoring systems is generally much lower compared to the three dimensional arrays of sensors deployed in the smaller and more complex deposits of A/C/S.

Figure 1. Inter-sensor spacing for A/C/S mines

Hudyma2008 estimated the inter-sensor spacing from several Canadian and Australian mines. This data from Hudyma2008 is displayed in Figure 1. The figure shows the cumulative distribution of the inter-sensor spacing with the markers scaled by monitoring area and marker colour showing the number of sensors used. The majority of the mines operate at a sensor spacing below 200 m. The mines with a larger sensor spacing are generally either very large

like the South African mines, or are not yet regarded as seismically active and only very few sensors were installed to assess the need for seismic monitoring. It should be mentioned that subsequent to the collection of this data, the general trend has been towards denser systems. The inter-sensors spacing in South African gold mines appears to be generally in the order of 300–650 m.

The tendency in the South African environment is to opt for triaxial sensors as this provides better quality data that facilitates source mechanism analysis. The A/C/S arrays tend to use a mixture of uniaxial and triaxial sensors with the aim at capturing smaller events and achieving a better source location.

Because these different approaches in the design of the seismic systems lead to differences in the data content, the approaches to extracting knowledge from these databases also differ. In South Africa, where the systems are generally less sensitive than in A/C/S, the approach is often to extract as much information as possible from the analysis of larger events, with the benefit of having more trained seismologists within the rock engineering community.

In A/C/S the problem of seismicity is generally the domain of rock engineers with limited formal seismology training. With the more sensitive systems in A/C/S, these rock engineers tend to focus on the overall rock mass response to mining, based on accurate source location and the analysis of populations of seismic events (clusters), including events as small as Richter magnitude -2.

Both approaches tend to converge into a somewhat similar risk management approach, where the estimation of the probability of seismic hazard is at the foundation of the different methodologies.

It is noted that the discussion put forward in this keynote has more relevance to the A/C/S approach of seismic monitoring.

SEISMIC SOURCES AND SEISMIC CLUSTERS

Seismic events tend to cluster at the locations of active seismic sources where some form of dynamic failure process occurs. The identification and understanding of seismic sources is important to seismic risk management in that they may be (or become in the future) the cause of significant seismic hazard. In particular, small events may start to form clusters at an early stage of extraction, with relatively small stress change. When the seismic system is sensitive enough to capture small events, this can assist in the timely identification of seismic sources and allow for the tracking of how these sources respond to mining, and more specifically, how seismic hazard related to these sources evolves as extraction progresses. Interpretation of the character of seismic sources in space and time is done by applying on individual clusters, series of analyses such as: magnitude-time history, the Guttenberg-Richter


relationship, apparent stress-time history, apparent volume, Omori relationship, and others (Hudyma2008).

Clustering seismic events according to seismic sources is at the foundation of the A/C/S seismic monitoring and risk management approach. The most widely used clustering method is probably the one implemented in the software MS-RAP based on the Comprehensive Seismic Event Clustering technique (CSEC) described by Hudyma2008. The CSEC method is a semi-automated, two-pass approach, which was developed on the basis of generic clustering techniques CLINK and SLINK (Jain et al.1999 and Romesburg2004). The first-pass of clustering using CLINK is totally automated, and groups together events exclusively based on spatial criteria. The CLINK clusters are then submitted to a second-pass of processing, where clusters are selectively grouped into “cluster groups” representing individual seismic sources. This cluster grouping is a manual process that requires interpretation of the likely seismic sources at the mine and requires a sound knowledge of the geology and the induced stress conditions prevailing at the mine. The cluster grouping is generally based on the similarity of source parameters, spatial proximity of clusters, and on the correlation of the location with known geological or geometrical features. The cluster grouping process can be seen as building a seismic source model by using the generated clusters as basic building blocks.

There are several practical difficulties associated with this process. The first relates to the definition of clusters on the basis of location only. The spatial coincidence of two different sources, say a fault in the vicinity of a raise experiencing dog-earing, will lead to the clustering together of fault events and the smaller events from the stress fracturing process. Another problem is related to the growth of clusters over time. In the first instance, a cluster is generated and grouped, and it is based on the events in the database at the time. As more data becomes available, this first clustering may need refinement and re-interpretation with obvious implications to the maintenance of the model of seismic sources.

Multi-dimensional clustering (i.e. basing the clustering on location and other source parameters) and grouping seems to be a logical solution but it is an open question if such an approach could be implemented and used practically.

As mentioned before, the grouping of clusters is in effect an exercise in building a seismic source model. In this sense it is similar to analysing an area with the use of polygons as the polygons become the basic units within the seismic source model. This process is, per se, subjective, and the value depends to a large degree on the understanding and training of the person performing this analysis.

This subjectiveness can to some extent be mitigated by performing spatial (grid based) quantitative analysis.

With this approach, the seismic source parameters are assessed through space by interpolating the source parameters. This approach allows for anomalies to be identified without prior selection of groups or polygons, as illustrated in Figure. 2 and 3.

Figure 2 shows the distribution of b-value in space at an Australian mine. A high value occurs around the stoping volume while a low b-value occurs at the lower abutment. These differences can be related to the difference in the source mechanism in these areas; the higher b-value corresponding with stress fracturing seismicity, and the lower b-value relating to a shear mechanism.

Figure 2. Grid based interpretation of seismicity distribution of b-value in space in a mining area at an Australian mine

Figure 3. Grid based interpretation of seismicity showing the distribution of b-value by colour and the size of the grid point marker scaled by apparent stress

This same plot is combined with a spatial distribution of apparent stress in Figure 3. The colouring of each grid point in space is the same as that in Figure 2 but in this figure each gridpoint marker is scaled by the geometric

Low Apparent stressHigh Apparent

stress

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mean of the apparent stress. The apparent stress is proportional to the mean shear stress at the source of the event (McGarr1994). The areas of the mine showing higher b-value generally show a low apparent stress while the lower abutment area shows a high apparent stress state corresponding with a low b-value. In this example, the distribution of b-value through space can be obtained without a pre-defined model of rock mass response. The use of spatial analysis of different source parameters provides a less subjective basis for building a model of seismic sources.

The use of these types of techniques requires high resolution data which can only be archived by dense array of sensors and high data quality.

ASSESSMENT OF SEISMIC RISK

The general definition of ‘risk’ is taken as the product of the hazard and its consequence. In terms of our discussion, which focusses on the risk to personnel rather than the more general risk to the mining assets, the consequence of the hazard is directly proportional to the personnel exposure and to the severity of the rockburst. In general terms thus:

Risk = hazard × severity x exposure

Assessing the seismic hazard is a non-trivial task and several different approaches exist. Approaches also differ for short, medium and long term.

Due to the short term volatility of seismicity, short term hazard assessments pose the biggest challenges. Generally short term hazard is assessed on the basis of studying time variations of a variety of seismic parameters effectively looking for ‘anomalies’ or precursory behaviour which are combined into a rating system.

Short term hazard fluctuations occur on top of what Wesseloo2013 referred to as the current hazard state. He assessed the current hazard state based on a spatial evaluation of the seismic density and frequency-magnitude distribution.

The general approach to medium and longer term hazard assessment is to estimate the magnitude of the largest probable event for a certain area. Such a design event will impact differently on different areas of the mine and is therefore scaled with a ground motion relationship to evaluate its effect at a distance. The peak particle velocity is generally used to quantify the effect of an event on excavations at a distance.

The hazard (PPV) map is an important step in assessing the seismic risk. The risk is simply related to the possible consequences of excavations being subjected to the ground motion (PPV) of the designed event (largest probable event). This may depend on the potential damage this event

can cause and the exposure of people and assets to the location of potential damage.

The above defines the variables controlling the seismic risk. If such risk becomes unacceptable, the mitigation can be achieved by reducing the hazard, reducing the damage potential, or reducing the exposure. These will be discussed in more detail in the following sections.

REDUCING SEISMIC HAZARD

A reduction of seismic hazard is generally not easily achieved. Some of the parameters influencing the seismic hazard at a mine, such as geology and ambient stress conditions, cannot be changed. The parameters that can be changed, like the mining method, mining sequence and mining geometry, require a significant amount of lead-up time to implement, as well as a large investment. These are sometimes referred to as strategic controls.

The theory underlying the reduction of seismic hazard is relatively simple and it is generally focussed on minimizing stress build-up on high hazard seismic sources. Numerical modelling techniques to assess stress redistribution in mines are widely used to investigate mining sequences that can minimize such stress build-up in seismically sensitive areas. These areas are often associated with seismically active geological features like faults, dykes, contacts zones, etc. or unfavourable mining geometries such as pillars, stope abutment, etc. The more severe sources of seismicity occur when poor mining geometry combines with adverse geological features.

Therefore, one way to reduce the seismic hazard is by using a more favourable mining geometry, and another is to create more favourable rockmass conditions.

Reducing Seismic Hazard with a more Favourable Mining Geometry

To influence the seismic hazard, these principles need to be translated into sound mine design practice. It is relevant to use some brief case studies to describe how seismic hazard has been reduced as a result of changes in the mining approaches at specific operations.

Mount Charlotte

Using the above principles and the benefit of their experience at the Mount Charlotte Mine, Mikula and Lee2000 developed a 14 point seismic risk reduction strategy for this Western Australian mine (Table I). These rules are based on the sound geotechnical principals of reducing induced shear stress on faults, reducing the build-up of high stresses on brittle pillars, and promoting non-violent deformation of rock mass closer to personnel occupied openings.

The development of this strategy enabled a systematic approach to the adaption of the mine design to reduce seismic hazard, and was very successfully used at Mount


Charlotte. All these rules may not be directly transferable to other mines. Providing mine design rules bases on sound geotechnical principles, which could be used by mine planners as a template for their mine design provides a good foundation for strategically managing seismic hazard.

Table I Seismic risk reduction strategy at the Mount Charlotte Mine

Seismic Strategy Item Summary

Backfill Use to add confinement, control loosening

Stiffness Keep local stope stiffness low, regional stiffness high, avoid irregular lumps and mining front. Faults reduce stiffness.

Access Do not mine accesses along faults. Avoid driving along any mining structure.

Unlock faults quickly Start mining at faults so it can move early. Intersect structure early.

West dipper Avoid undercutting or overcutting west dipping structure.

Stope end abutment access Orient abutment development E–W.

Stress shadow the faults Unclamp faults and encourage gradual movement.

Abutment stress Use narrow E–W abutments. Avoid increased shear stress on clamped faults.

Pillars Avoid diminishing pillars. Avoid stress increases in and around pillars.

Blasts Small blasts generally cause small stress change, and are associated with small energy releases.

Preconditioning Intentionally weaken ground to reduce its ability to carry stress, and encourage movement.

Destressing Redirect locally high stresses.

Blast timing Die-down exclusion time of two hours.

Reoccurrence New events are less likely in ground distressed by previous large events.

Laronde Mine

Mercier-Langevin2008 describes the Laronde mine undergoing successive expansions from 980 to 2240 m below surface using a bottom-up chevron-type sequence (Figure 4a) where the front of stopes ‘plough’ through the highly stressed sill pillars (Figure 4b). Experience at this mine has shown that an area of high seismic hazard always developed on the east (right) side of the chevron as the mining front breaks through the upper levels (Figure 4b, area defined by the oval). Areas with a high yearly probability of a seismic event exceeding ML 2 shows up as solid in the figure. In Figure 5, a new sequence devised to mitigate the seismic risk on the eastern flank below the 2240 horizon is shown.

Mercier-Langevin2008 described the sequence as follows: “After the 4 initial stopes, the western flank is mined as per

the standard sequencing (primary and secondary stopes). The primary stopes on the eastern flank however are mined in an underhand fashion”.

Figure 4. The stacking “chevron” mining sequence adopted at Laronde (a). Probabilistic hazard assessment showing high hazard to the east of the chevron as the mining front breaks through (b)

The top-down sequencing of stopes on the eastern side of the chevron (Figure 5) results in the early extraction of the highly stressed sill pillar on the seismically sensitive east side, and also the elimination of the sharp corner which previously concentrated the stress.

With the modified sequence, high stress concentration on the seismically active east side was avoided due to the early extraction of sill pillar on that side and also the more favourable obtuse angle geometry resulting from the top-down mining sequence.

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Figure 5. The modified chevron, using a top-down sequence on the eastern front, creating an obtuse angle to avoid stress concentration on the seismically active side

Junction

A somewhat similar strategy is described in Li et al.2007 where the mining method and sequence at the Junction mine was totally reworked to reduce the seismic hazard which had become unacceptable as the extraction progressed deeper.

Li et al.2007 described the original mining method as follows: “...a longhole retreat method, with pillars and delayed unconsolidated fill. The mining sequence was typical in Western Australia, that is, to retreat from either extremities of the orebody to a central pillar”. This indeed resulted in a high stress concentration and seismic response in the diminishing pillars. This “typical” Western Australian mining sequence is particularly attractive economically, but becomes impracticable at depth, as discussed in Turner and Beck2007.

At about 400 m below surface the method was modified. Still using unconsolidated fill, the pillar size was increased, the strike length of stopes was reduced and the retreat sequence was changed to eliminate the “shrinking pillar”, progressing from one end of the orebody to the other. This new approach improved results down to

approximately 600 m below surface. Nevertheless, despite these changes in the mining method, it was assessed that excessive dilution and rockbursts would make the deeper levels unrecoverable.

A completely new method was conceived to control dilution and seismic hazard. The new sequence used a continuous “pillar-less” retreat, progressing top-down in an inverse chevron shape (Figure 6). High cement content paste fill was introduced to allow mining to take place top-down, directly under filled stopes. The process of implementing the new mining method at Junction took 19 months, from the evaluation of mining sequences which started in September 1999 to the extraction of the first stope in April 2001. This strategic change reduced the seismic hazard and allowed the full and successful extraction of the orebody at depth.

These case studies have shown that it is possible to reduce seismic hazard though a strategic approach, but that it requires significant forward planning and investment. It is noted that the Big Bell mine and the Tasmania mine had to go through similar process following their fatal rockburst accidents.

Although simple in theory, the execution of some of these strategies can sometimes be challenging. The ideal mining sequence for seismic risk reduction is rarely the most attractive, economically. Then compromises must be assessed to deliver maximum value to the mine with an acceptable risk. Mine design must therefore be performed on a risk reward optimization with due attention to safety, as has been argued for open pit design by Wesseloo and Read2009 and Stacey et al.2006 for underground rock engineering design in general. Currently however, clear guidelines and tested and trusted methodologies for performing mine design from a risk basis for different mining environments does not exist.

Figure 6. Longitudinal section of Junction Mine showing the new direction of retreat, pushing stresses towards the abutments


Reducing Seismic Hazard by Changing the Rockmass Conditions

Another approach aimed at reducing the seismic hazard is to induce fractures in the rock mass and change its seismic response to mining. The ‘pre-conditioning’ technique can be done by hydraulic fracturing or de-stress blasting. Although not entirely understood, it is believed that pre-conditioning changes the capacity of the rock mass to store strain energy.

De-stress blasting, at the scale of a drive, has been practiced in mining for many decades, but it is still considered more an art than a science. At the development scale, the seismic hazard is often relatively low, but the exposure of the workforce is very high. In this case the aim is to influence only the local environment around the excavation to reduce strainburst failures.

Larger scale de-stress blasting of an entire stope or pillar have also been attempted occasionally (Andrieux et al.2010). At an even larger scale in cave mines, it has been reported that systematic hydro-fracturing of the entire caving column has resulted in a significant reduction in large seismic events at the Codelco El-Teniente Mine (Araneda and Sougarret2007). These relatively new approaches have the potential to reduce seismic hazard, without having to re-engineer the mining method.

More research and development is required to improve our understanding, and the limitation, of these techniques.

REDUCING DAMAGE POTENTIAL

Notwithstanding the relatively new large scale de-stressing options, once the mining method and sequence are set, the seismic hazard state, to a certain extent, is more or less “locked in” for the foreseeable future. The only short term controls (possibly apart from the pre-conditioning and de-stress blasting discussed above) are to reduce the consequence parameters of the seismic risk equation, i.e. damage potential and exposure of people and assets.

The potential rockburst damage is arguably best mitigated with the implementation of dynamically resistant support. New bolts and new surface support methods capable of withstanding large deformation and significant loads have been developed in recent years. The laboratory performance of many of these new products are reported in the proceedings of the Deep and High Stress Mining Seminar series (Potvin2007, Van Sint Jan and Potvin2010, Potvin2012) as well as in the 6th International Symposium on Ground Support in Mining and Civil Construction (Stacey and Malan2008). In addition to these tests, anecdotal evidence has shown that some of these dynamically resistant systems can be quite effective in reducing rockburst damage. However, there is a perception that they are expensive, especially when they are compared to the cost of conventional support. This is indeed ignoring

the value of risk reduction. Nevertheless, only few mines to date have chosen to implement systematic dynamically resistant support throughout their operations. Most mines instead only use dynamic resistant support in areas previously identified as being exposed to elevated seismic hazard, using hazard maps or similar techniques.

One of the current challenges in terms of dynamically resistant support lies in the engineering design aspect. In countries where seismic hazard is significant, regulators will demand that a proper engineering design for the dynamic support system be prepared. This is in spite of the fact that current methods are inadequate to complete this task. Stacey2012 argued that the design of dynamic support is clearly a case of “design indeterminacy” because there are no means of reliably estimating the demand or the capacity of ground support subjected to the dynamic loading of an incumbent seismic wave (Potvin and Wesseloo2013).

There has been significant recent research effort aimed at improving this situation. In particular, the dynamic capacity of rockbolts has been investigated with a number of “drop testing rigs” and reported in Yi and Kaiser1994, Kaiser et al.1996, Ortlepp and Stacey1997, 1998, Ortlepp et al.1999, Stacey and Ortlepp1999, Ortlepp and Swart2002, Gaudreau et al.2004, Plouffe et al.2008, Player et al.2004, 2008a,

2008b, 2009, Villaescusa et al.2005 and Tannant et al.1993, 1994. Potvin et al.2010 compiled the results from many of these tests, showing the trends between energy absorption and displacement of reinforcement and surface support (Figure 7).

Despite the differences in the testing methodologies used by different researchers, the trends are surprisingly consistent. The light green shaded area (on top) describes the approximate behaviour of grouted reinforcement (including de-bonded bolts, but excluding friction bolts split sets), the lower blue shaded area represents fibrecrete or welded mesh commonly used in mines, and the middle purple zone characterizes high capacity surface support. It should be noted that the high capacity surface support was obtained through the use of cable or steel rope lacing.

The drop-testing results are very useful to investigate the behaviour of ground support under dynamic loading and establish comparisons between different products. However, the capacity obtained from the tests cannot be used directly for design as the loading mechanisms and the load transfer mechanism is not similar to in situ conditions. The extrapolation laws, from laboratory tests to field performance, have yet to be established in this case. As a consequence these test results should be seen as index values of the relative strength.

The calculation of a realistic dynamic demand on ground support is perhaps even more difficult to assess. Kaiser2013 rightly makes the distinction between a “designed dynamic demand” value and the real demand, the

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Figure 7. Total energy absorption and displacement of reinforcement and surface support

dynamically loaded using drop weight tests

designed being at the upper bound of the real value. From experience and intuition, the real dynamic demand on ground support should be much greater for large seismic events at close proximity to an excavation than small events located remote from the excavation. But there are many exceptions to this rule, as sometimes large events close to excavations do not produce any damage and relatively small events relatively far away trigger significant damage. This is believed to be due to the many factors that may influence the outcome of a seismic event. Heal2010 suggested that the main factors are:

Strong ground motion.

Induced “static” stress around the excavation prior to the event.

Rock mass behaviour (brittleness, quality, presence of faults, etc.).

Size of the excavation.

Capacity of the ground support system.

Generally the peak particle velocity (PPV) is used as a parameter to quantify the strong ground motion. Notwithstanding the last four parameters, estimating the PPV at excavation surfaces alone is a difficult challenge. It is sometimes questioned whether PPV is an adequate measure of the effect of the incumbent stress wave. Modern seismic systems can locate and produce a measurement of the intensity (magnitude) of seismic events at the source. To transform this measurement into PPV at the excavation

surface, one has to consider the radiation pattern and the attenuation of the stress wave as it travels through the rock mass. It is known that the propagation pattern and the attenuation are likely to be very complicated and difficult to model.

Furthermore, Milev et al.1999, Durheim et al.1998, Hagan et al.1999 and Durheim2012 describe “a site effect” which amplifies the stress wave as it transits from a confined environment within the rock mass to the unconfined excavation surface. Accounting for the site effect adds a further complication in the estimation of the resulting ground velocity at the surface of an excavation.

Finally, the degree of rock brittleness may also have a strong influence on the amount of energy that can be stored in the rock mass, and later transformed into kinetic energy during the failure process, as described by (Tarasov2010, 2011, Tarasov and Randolph2011, Tarasov and Potvin2012).

The above factors are discussed in more detail in Potvin and Wesseloo2013, but no clear solutions are offered to estimate the dynamic demand on ground support. It is the authors’ opinion that much research needs to be done before reliable methods to design dynamic support become available. Until science progresses sufficiently to provide a reliable deterministic approach based on capacity and demand, perhaps empirical methods, such as the ones proposed by Heal2010 and Mikula2012, present an alternative approach to assess rockburst damage potential.


REDUCING THE EXPOSURE

Reducing the exposure of assets and, more specifically, people since they are the most valuable asset is the last line of defence against seismic risk. The strategic decisions taken long before extraction begins will have a significant impact of the exposure of people and therefore on the seismic risk at the mine. For example, the selection of highly mechanised mining methods with low exposure of people to production areas where frequent stress change occurs, and the degree of automation implemented to remove workers from hazard zones, are very effective ways of reducing seismic risk. However, the risk can never be completely eliminated.

The “Holy Grail” of mine seismicity is the prediction of the time-location-magnitude of large events based on precursors. If this could be achieved, it would allow for the evacuation of the areas at risk in time to avoid exposure of people to rockburst damage. Unfortunately, the current state of knowledge is not even close to achieving this. Spottiswoode2010 has questioned the notion that seismic precursors exist on which such decisions can be made. Van Aswegen2003, 2005, however, claims some success and has quantified the success rate of the short term hazard prediction (van Aswegen2013). He also showed that the success rate varies for different conditions and that under some conditions a reasonable success rate can be achieved. It would be safe to say that precursors of large events remain poorly understood and in recent times, the emphasis of researchers has been concentrating on other risk management techniques, as described above. Even if significant progress was achieved in this area, it is questionable whether such an approach could be implemented effectively in A/C/S mines. The liability of mine management and technical experts responsible to ‘set the alarm’ would probably be untenable. A parallel can be drawn here to the six Italian scientists who were found guilty of multiple manslaughters for underestimating the risks of the L’Aquila’s earthquake and failing to alert the population (BBC2012). Notwithstanding the above comment, the inclusion of short term seismic hazard assessment as a part of the overall seismic risk management strategy (and not as the prime mitigation tool) would be highly desirable, provided that the reliability of the technique used is acceptable.

Reducing the exposure of people can also be achieved through applying “re-entry” rules. A certain proportion of seismic activity and large events occurs as a direct response to blasting. The seismic response can be “gentle” or “violent” and it can occur shortly or a long time after blasting. In seismically active mines, a decision must be taken after each blast (almost daily), as to whether it is safe for the workforce to return to their respective workplaces. To reduce the exposure of people to these blast related seismic events and aftershocks, it is common for mines to

develop re-entry rules and protocols based on the study of after blast temporal decay of seismicity, amongst other factors. It is generally accepted that when the seismicity returns to (or near to) the background level, then the stress has been redistributed and has reached some form of equilibrium, the seismic response is completed, and the hazard has reduced to ambient levels. Other geotechnical monitoring techniques and engineering judgement can also be used to increase confidence in the decision making process.

Examples on how this kind of approach has been applied in mines are given in Heal2007 and Hudyma2008. To analyse seismic decay, some parameters must be arbitrarily defined, i.e. the volume of influence from the blast, the size of the time window after blasts, and the minimum total seismic energy (because aftershock sequences with limited combined radiated energy is not regarded as a risk to cause damage). The seismic decay is assessed for the event rate (most common), seismic energy, seismic moment, the percentage of decay which is considered sufficiently close to background level for re-entry. These approaches however, are somewhat convoluted and lack theoretical and statistical rigour and cannot be applied on a real-time basis.

Recent work by Vallejos and McKinnon2008, 2010, 2011 has provided a more rigorous approach based on statistical characterisation of seismic decays to develop protocols to assess safe re-entry time after blasting. They also provided some methods for evaluation of the seismic decay for real-time re-entry evaluation.

The approaches mentioned above, are “blast focussed”. Woodward2013 argues for an adjustment in the paradigm from ‘blast focussed’ to “rock mass focussed”, in other words not asking, “How did this blast affect seismicity?” but rather “How did this volume of rock mass respond to the blasting?” This approach will enable a characterisation of the way different rock masses react to blasting in different areas, which will lead to better control over exclusion zones.

An example of this is provided by Woodward2013. He showed a very strong response of a particular area in a mine to blasting up to 250 m away. This is illustrated in Figure 8, showing the blasts as stars and the mean location of a ‘seismic sequence’ with X’s (a "seismic sequence" describes a flurry of events in time and space). The figure shows only those seismic sequences that reacted immediately to the blast but were remote from the blast. This particular area is sensitive to static and/or dynamic transient stress changes caused by blasting some distance away.

This has important implications for developing re-entry protocols, as exclusion zones should not only take into account where the blast is taking place, but also how different volumes of the rockmass react to it.

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Figure 8. Blasting and the centre location of remote “seismic sequences” associated with the blast. The contact between two different rock masses are also shown. Sequences locate with a stiff brittle highly stressed rock mass

THE TASMANIA MINE; A CASE STUDY OF SEISMIC RISK MANAGEMENT

The Tasmania mine is an excellent case study to illustrate many seismic risk management concepts described in this keynote. The authors have sourced most of the information in this section from (Hills2012, Goddard and Hills2011, Hills et al.2008).

The Tasmania mine is essentially a shear hosted vein that could be described as a tabular narrow vein orebody dipping at 50–70 degrees to the south-east. Figure 9 is a longitudinal section showing the mine workings in relation to the stratigraphy of the ore deposit.

The Tasmania mine had two main mining periods. The original mine started in 1877 and the extraction progressed from the surface to 455 m below surface. The mine was shut down in 1914 and remained dormant for a long time. No issues with seismicity were reported in this first and shallow phase of extraction. The second phase of mining commenced when production resumed in 1999, following a three year feasibility study. The extraction progressed from 455 m to 1210 m Level and was completed in June 2012. From the feasibility work, it was anticipated that the onset of seismicity would occur below the 700 m level, and this was proven correct. In particular, the western side of the mine responded seismically to mining activities. The Tasmania mine was pro-active in the management of seismicity from that point in time, which was around 2002–2003. Hills2012 reported that:

“The local stress regime was measured, understood and modelled in 2003. A temporary seismic monitoring system was operational from April 2004 and it was replaced with a permanent array in mid 2005”.

At that time, the Tasmania mine was also amongst the few mines to implement dynamic reinforcement in Australia, being the first to adopt the modified cone bolts in an area where elevated seismic hazard was anticipated. Strategic changes that were implemented to lower the seismic hazard included:

Widening of sill pillars to improve their load bearing capacity.

The change in the mining retreat sequence to maintain a steeper advancing front.

Regular peer reviews of the mining and geotechnical plans.

Adoption of an after blast exclusion period of 24 hours to reduce the exposure of the workforce to blasting induced seismic aftershocks.

As mentioned earlier, in seismically active mines the risk cannot be entirely eliminated, and despite all the effort to manage seismicity, Hills2012 concluded that:

“It wasn’t enough: a seismically induced rockfall on 25th April (Anzac Day) 2006 took the life of a colleague”.

Without going into detail of the accident, the fatal rockfall occurred in the ore-drive of the 925 Level as a result of seismic shakedown. A further large shakedown occurred simultaneously in the ore–drive of the 915 Level. Ore-drives in previously extracted levels had been hit with seismic damage in the past and were experiencing elevated seismic hazard, which was controlled with dynamic ground support. Therefore, one of the main strategic risk mitigation measures after the accident was to re-engineer the mining method to completely eliminate the exposure of mine workers to the ore–drives exhibiting elevated seismic hazard. As at the time of the accident some of the ore-drives were already developed, an innovative method was devised on the basis that no activities were to occur in theore–drives. This was named the “Western Method Stoping”.

The method is discussed in detail in Hills et al.2008 and summarised in Hills2012 as follows:

1) Every existing sill drive was duplicated by a footwall drive at a minimum stand-off distance of 6 m.

2) The footwall drive was reinforced with a dynamic upgrade such that it could withstand a seismic event of maximum magnitude ML 2.5 at a distance of 10 m with a factor of safety of 1.5.

3) A nominal 20 m strike length stope (approximately 2000 t) was drilled using rings of fanned holes from the footwall drive.

4) The stope was blasted in a horizontal chevron pattern utilising the existing sill drive as the free face using rapid electronic initiation. No slot rise was required.


Figure 9. Longitudinal section of the Tasmania Mine, showing the relative location of the mine workings with the stratigraphy

5) The stope was mucked using tele-remote controlled loaders.

6) The resulting excavation was surveyed using a remotely operated laser scanner.

7) The stope was backfilled from the sill drive above using tele-remote loaders.

8) The process was repeated for the next stope, sequenced using the checkerboard initiative.

The Western Method stoping arrangement is illustrated in Figure 10a.

The Western Method stoping further evolved to the more operationally and financially attractive Radial-in-Reef stoping (Goddard and Hills2011), which used full ring drilling from the footwall drive and allowed for the extraction of full height stopes, resulting in less development mining per ton extracted. It is noted that an Avoca type of mining where ore-drives are accessed by mine workers was also practiced in areas where the seismic hazard was low. Having a good understanding of the seismic hazard throughout the mine remains one the keys to a safe operation. Seismic hazard maps such as the one shown in Figure 11 were kept up-to-date since production resumed at the Tasmania Mine in 2007 after the 2006 rockfall.

The dynamically resistant ground support continued to evolve and was improved to further reduce the rockburst damage potential. A system similar to the one implemented in the deep nickel mines of Sudbury, Canada was adopted in the elevated hazard areas of the Tasmania mine, relying on de-bonded cable bolts and mesh straps to increase the energy absorption capability of the system.

It was mentioned earlier that the identification and tracking of seismic hazard is at the foundation of seismic risk management. To assist in the interpretation and tracking of the seismic hazard, one of the most dense instrumentation programmes was implemented at the mine. More than 200 individual instruments were deployed in 80 clusters.

Figure 12 shows the location of the stress measurement devices at the Tasmania mine. Apart from these devices, many extensometers and/or SMART cables were installed.

This was used as an integrated approach to geomechanics and seismicity management applied to all aspects of mining. The exclusion and re-entry protocols after blasting and large seismic events were managed based on the better understanding of the seismic response achieved with conventional and seismic monitoring equipment, thus further reducing the exposure of mine workers. Also the instrumentation was useful in assessing

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Figure 10. Drawings illustrating a) Western Method

stoping, and b) Radial-in-Reef stoping

Figure 11. Deterministic seismic hazard map at the Tasmania mine

Figure 12. Stress measurement locations

at the Tasmania mine

the rock mass response to the numerous changes made to the mining methods and production techniques.

Figure 13 (Hills2013) shows the strong correlation between the blasting and the stress change measured by instruments close by. Similar stress change was measured following large seismic events in the vicinity of stress measurement devices. We believe that using such an approach of combining seismic and non-seismic data to build a more complete understanding of the rockmass response, confirming design assumptions and, providing more confidence in the interpretations, has a great potential to improve seismic risk management.

This case study has shown how seismic management is a complex process. As confirmed by the Beaconsfield Coronial Investigation, prior to 2006, an adequate seismic management system was in place in accordance with world good practice. Attempts were made to reduce the hazard through changes in the mining method. The reduction of the damage potential was also addressed by installing the modified cone bolts, a dynamic resistant support. A seismic system was installed to understand and monitor the seismic hazard, and after blasting exclusion and re-entry rules were developed.

Following the 2006 accident, the seismic risk was further lowered significantly by eliminating the exposure of mine workers to high hazard ore-drives, with new and innovative mining methods. The ground support system was up-graded with increased energy absorption capacity


Figure 13. Illustration of measured strain gauge and interpreted stress change resulting from blasting (Hills2013)

and the exclusion re-entry protocols were managed based on one of the most comprehensive instrumentation programmes ever installed for the size of the mine.

It would be safe to say that the instrumentation programme at Tasmania mine was unique in its extent. The type of integrated interpretation of seismic and non-seismic data employed at the Tasmania Mine is not common. In our opinion, only a fraction of the value was extracted from the data collected from the Tasmania mine. This was mainly because methods and systems need to be developed to extract information efficiently in a practical

timeframe. The development of integrated analysis systems is an important area for future development.

CONCLUSION

The mining industry has come a long way in managing the risks associated with mine seismicity and the statistics from many countries around the world show a consistent downward trend in rockburst fatalities, despite deeper mines and higher stress conditions.

The risk mitigation based approach relying on understanding and reducing seismic hazard and people

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exposure with more appropriate mining methods and extraction sequences, and reducing the potential consequences of large seismic events by limiting damage potential with improved ground support systems and design techniques, has proven to be very effective. This keynote is aimed at providing details and examples of the seismic risk based approach, and outlines some of the current limitations to assist in focussing future work.

It is the author’s view that it is important that all aspects of the management of seismic risk are addressed in on-going research and development, and special attention should be given to the following areas:

Improving the understanding of the mechanism and quantifying the influence of preconditioning on the seismic hazard.

Improving the understanding and quantifying the ground-support interaction under dynamic loading.

Improving the understanding of the ground response to blasting.

Further develop the methods for integrated interpretation of seismic and non-seismic data.

ACKNOWLEDGEMENTS

The ACG sincerely thanks the following organisations who provided funding for this research through the Mine Seismicity and Rockburst Risk Management project: Barrick Gold of Australia, BHP Billiton Nickel West, BHP Billiton Olympic Dam, Independence Group (Lightning Nickel), LKAB, Perilya Limited (Broken Hill Mine), Vale Inc., Agnico-Eagle Canada, Gold Fields St Ives Gold Operations, Hecla USA, Kirkland Lake Gold, MMG Golden Grove, Newcrest Cadia Valley Operations, Newmont Asia Pacific, Xstrata Copper (Kidd Mine), Xstrata Nickel Rim, The Minerals and Energy Research Institute of Western Australia. We also thank Peter Hills for his comments on the original manuscript.

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Keynote Lecture: Improving seismic risk management in hardrock mines

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