Difference between revisions of "Factors influencing mining method selection"
(→Geotechnical Factors of Underground Mining Methods)
|Line 168:||Line 168:|
== Summary ==
== Summary ==
A summary of the geotechnical factors for each underground mining method, including the suitable orebody geometries, orebody grades, orebody and country rock strengths, and depths are shown in Table 1.
Revision as of 14:31, 25 July 2012
From Queen's Mine Design Wiki
Note: Oil and gas deposits are not discussed in this article. Underground mining methods are the focus of this article.
The selection of a mining method for an ore deposit is based on many factors that are driven by the economics and profitability of the mine for a company. These include the ore grade and recovery, cost of infrastructure, ore extraction, labour and machine costs, and underground support costs. The characteristics of the orebody itself form the basis for these decisions, including the thickness and orientation of the mineralization, the ore and rock strength, the distribution of mineralization within the orebody, the geotechnical environment, and the depth of mineralization and surface conditions. In some cases, these conditions change in a single mining operation. If significant enough, a change in mining method in one ore deposit can occur.
Geotechnical considerations when selecting a mining method are a relatively recent trend, caused by the increased dimensions and production rates required of mining operations in order to meet growing expectations of profitability. Since these larger projects require a longer period of satisfactory performance in terms of ore recovery and ground support, more formal and rigorous methodologies are necessary in mine design (Brady and Brown, 2006). Geotechnical factors include in-situ mechanical properties of the orebody and country rocks, the geological structure of the rockmass, the ambient state of stress and the hydrogeological considerations in the zone of potential mining influence (Brady and Brown, 2006). The goals of geotechnical consideration in mine design, regardless of the mining method, are to:
- Ensure the overall stability of the complete mine structure, defined by the main orebody, mined voids, ore remnants (pillars) and adjacent country rock;
- To protect the major service openings and infrastructure throughout their design life;
- To provide safe access and working places in and around the centres of ore production; and
- To preserve the mineable condition of unmined ore reserves (Brady and Brown, 2006).
Mining methods have evolved significantly in the last several decades as improvements have been made on machinery used to extract the ore, understanding and experience with the behaviour of the rockmass and underground stresses has developed, and as newly discovered ore bodies are located in increasingly difficult conditions.
- 1 Mining Method Classification
- 2 Change of Mining Method in a Mining Operation
- 3 Thickness and Orientation of Mineralization
- 4 Ore and Country Rock Strength
- 5 Distribution of Mineralization within the Orebody
- 6 Depth of Mineralization and Surface Conditions
- 7 Geotechnical Environment
- 8 Geotechnical Factors of Underground Mining Methods
- 8.1 Pillar Supported
- 8.2 Artificially Supported
- 8.3 Unsupported
- 9 Summary
- 10 References
Mining Method Classification
From a geomechanical perspective, mining methods can be classified based on the type and degree of support required in mining operations. Supported mining methods include open stoping and room-and-pillar mining, where natural support is provided by ore remnants (e.g. pillars), or cut-and-fill stoping and shrinkage stoping, where support for the walls of the void remaining after ore extraction is provided by backfill or by fractured ore temporarily retained in contact with mined stope walls. Cave mining methods include block caving and sublevel caving, where no support is used because fragmented rock fills and flows through the stopes. A classification of underground mining methods, subdivided based on pillar supported and unsupported groups, is shown in Figure 1.
The distinction between these two broad categories of mining methods can be made by comparing the displacements induced in the country rock and energy redistributions in the rockmass caused by mining activities. Supported mining aims to restrict displacements in the country rock to elastic behaviour and prevent failure of the rockmass. The success of these methods depends on the ability of the near-field rockmass to sustain compressive stresses in order to maintain elastic behaviour. The mining issue therefore becomes prevention of unstable energy releases (e.g. rockbursts) associated with increased near-field stress, which could cause failure of support elements, sudden closure of stopes, or rapid fracture generation in the surrounding rock. A schematic of a supported mining method (room-and-pillar) is shown in Figure 2.
On the other hand, cave mining purposefully induces large displacements following fragmentation of the rockmass, resulting in energy dissipation in the caving rockmass. The success of this method depends on exploiting the discontinuous behaviour of a rockmass when confining stresses are relaxed (Brady and Brown, 2006). The mining issue here is to maintain steady displacement of the fragmented orebody so to prevent the development of unstable voids. The rate of slip and fragmentation of the rockmass must be proportional to the rate of ore extraction. A schematic of a caving mining method (block caving) is shown in Figure 3.
Change of Mining Method in a Mining Operation
In practice, it is possible for a mining operation to utilize different mining methods, which can even be classified by different geomechanical concepts, at different stages of orebody extraction. The transition between methods of different geomechanical behaviour can have significant consequence on the stability of permanent openings, which lends to the importance of defining a mine plan for the entire mine life that will be able to successfully accommodate the necessary mining methods and induced behaviours of the rockmass. However, in underground mines, a complete change of mining method once operations have begun is uncommon. Slight variations in the method occur when faced with minor changes to the ore body and mining environment, but a complete overhaul of the initial method requires a change that significantly impacts the mine output and its overall profitability. In general, mines that have experienced continuous problems are more willing to adopt new mining techniques to improve their operations with a changing mining situation (Laubscher, 1994).
Thickness and Orientation of Mineralization
Orebodies can occur in a variety of geometries, related to the deposit’s geological origin. Tabular or stratabound deposits are of sedimentary origin and are extensive in two dimensions (horizontally if in unmetamorphosed sedimentary rock). Veins, lenses and lode deposits are also generally extensive in two directions but are formed by hydrothermal and/or metamorphic processes. Massive deposits have a more regular orebody shape that are controlled less by geologically imposed boundaries (Brady and Brown, 2006). The details of preferred orebody shapes are discussed in the section for each mining method. A detailed discussion of oil and gas deposits is not included in this article.
Ore and Country Rock Strength
The strength of the ore and the surrounding rock is one of the most significant geotechnical factors in mining method selection. The strength and related stability of the rockmasses are important for all types of excavation, and are assessed using rockmass classification systems throughout a mine. Common rockmass classification systems include Bieniawski’s Rock Mass Rating (RMR) system (Bieniawski, 1989), the Modified Rock Mass Rating (MRMR) system developed by Laubscher (1990), the Q system developed by Barton (Barton et al., 1974), and the Geological Strength Index (GSI) by Hoek and Brown (1997). Support requirements within the guidelines of a mining method are determined based on the strength of the rockmass as well as the intended use of the excavation (e.g. permanent service excavations versus stopes) and the related risk of failure to the mine.
The RMR system was developed to assess the stability of an excavation, while the MRMR system addresses cavability of an orebody being mined using caving methods. Although the MRMR system has been used successfully for the weaker and larger orebodies for which it was first developed, more recent experience in stronger, smaller and isolated or constrained orebodies has not provided satisfactory results (Brown, 2003).
The mineralization in a rockmass that defines the orebody changes the geomechanical properties of that rockmass. The intact rock strength as well as the amount and quality of joint sets and fractures of an ore body are usually different from the surrounding rock. The type and grade of mineralization affects the sharpness of the contact between the rock types, which is an important control on dilution during mining.
Distribution of Mineralization within the Orebody
The variation of ore grade through the volume of an orebody influences the mining strategy. The critical parameters are average grade, cut-off grades, and grade distribution. The average grade determines the degree of flexibility for method selection as related to the operating costs and current market conditions that define the monetary value of the deposit. The amount of dilution of ore expected during extraction is also related to the value per unit weight of ore. For deposits with a lower average grade, there is a higher economic sensitivity to the effects of dilution.
General grade distribution in an orebody may be uniform, uniformly variable, or irregular. Uniform ore grades are found in massive ore deposits, uniformly variable grades exhibit a spatial trend in ore grade, and irregular ore grades are found in deposits with high local concentrations of ore minerals, including vein, lens, and nugget deposits.
Depth of Mineralization and Surface Conditions
The geotechnical environment of an ore deposit is characterized by the intact rock and rockmass properties, in-situ stress in the host rock, and chemical properties of the ore (Brady and Brown, 2006). Intact rock properties include strength, deformation characteristics and weathering characteristics. Rockmass properties are defined by the influence of joint sets, faults, shear zones, and other penetrative discontinuities.
Adverse chemical properties of ore rock may prohibit caving methods of mining, which generally require the ore to be chemically inert (Brady and Brown, 2006). For example, a sulphide ore subject to rapid oxidation by fragmentation of the orebody may create difficult ventilation conditions in working areas and even break down into smaller pieces of rock after predicted primary and secondary fragmentation have occurred. These smaller rock fragments could reduce the effectiveness of the height of draw in the stope and the transport and handling facilities for the ore.
There are certain cases where a pervasive geological feature can be influential enough to control the entire mining method selection and mine plan, including large fault or shear zone systems, highly fractured rock linking to an aquifer, and the local tectonic setting (Brady and Brown, 2006). Faults and shear zones may separate the orebody into multiple sections that would otherwise be able to be mined using large scale caving. Similarly, aquifers existing in or near the zone of mining influence where large fractures occur in the rockmass may provide hydraulic connections to other water sources during mining. An active tectonic setting would be problematic for large voids left by mined stopes, due to the possibility of local instability induced by a seismic event. Also, mining activity in a stope near an existing fault has the potential to cause a local seismic event. In addition to the direct dangers of structural failure of the rockmass in the stope, an indirectly caused an air blast is a consequential risk for mine safety.
Geotechnical Factors of Underground Mining Methods
The discussion of specific underground mining methods is organized based on the type and degree of support required in mining operations: pillar supported, artificially supported, and unsupported. Case studies are presented on mining methods at the Kristeneberg Mine (cut-and-fill) and the Mouska Gold Mine (shrink stoping).
The successful performance of a pillar supported system is related to both the dimensions of the individual pillars and their geometric location in the orebody (Brady and Brown, 2006). A good understanding of in-situ stress conditions is necessary for a successful pillar supported mine design. If there is a high horizontal maximum stress in a particular direction, the orientation of both the room advance through the orebody and rectangular pillars should be planned in order to maximize support in that direction. Many very shallow room-and-pillar operations may have very little horizontal stress such that the orientation of the rooms and pillars has a minimal effect. However, very deep operations in high stress environments may have rockburst issues. As such, the sequence of extraction in addition to pillar orientation is important (Bullock and Hustrulid, 2001).
Room and Pillar Mining
Room and pillar mining generates ore pillars as remnants as extraction progresses, in order to control the stability of the roof rock and the global response of the surrounding rockmass. Regular patterns of pillars are typically developed in order to simplify planning, design, and operation. The roof may or may not be artificially supported for worker safety, depending on the competency of the rockmass. Pillars can either remain intact upon mine closure or extracted at the end of mine life, allowing the stope to collapse afterward (Brady and Brown, 2006).
Room and pillar mining is best suited for tabular deposits that also must be relatively shallow to limit the size of the ore pillars. Examples of tabular ore bodies or host rocks include copper shale, coal, salt and potash, limestone, and dolomite (Hamrin, 2001). There are three typical variations of room and pillar mining that account for changes in dip of the ore body.
- Classic room and pillar mining is used for horizontal deposits with moderate to thick beds, as well as inclined deposits with thicker beds. Mining progresses downward from the hangingwall in slices and the required ground support is installed in the hangingwall.
- Post room and pillar mining is used for thick, inclined deposits that have a dip between 20º and 55º. The mining sequence here begins from the bottom and advances upward. Backfill is used to increase the support capacity of the pillars in the mined out areas, and to create a platform from which to mine the next section of ore.
- Step room and pillar mining is used for ore deposits that are only 2 to 5 m thick and have a dip angle between 15º and 30º. In this case, mining advances downward from the hanging wall (Hamrin, 2001).
A suitable geomechanical setting for room and pillar mining requires a strong, competent orebody and near-field rockmass, with a low frequency of cross jointing in the immediate roof rockmass (Brady and Brown, 2006).
Mississippi Potash Inc.’s underground operations are a good example of room and pillar mining of soft rock. The salt ore is surrounded and intruded by clay seams which form zones of weakness that are controlled by rock bolts or cribs. Since the stability of the salt layers is much easier to control with ground support, the roof of the mine excavations are designed to be developed in salt (Herne and McGuire, 2001).
Sublevel Open Stoping
Sublevel open stoping requires extensive development in and around the orebody during preproduction. Stope faces and side walls remain unsupported during ore extraction, while support for the country rock is developed as pillars are generated by stoping (Brady and Brown, 2006). The pillars may be left in place or extracted at a later time (Bullock and Hustrulid, 2001). Bighole stoping is a larger scale variant of sublevel open stoping that uses longer blast holes. This results in vertical spacings between sublevels of up to 60 m instead of 40 m for sublevel open stoping. Sublevel open stoping is applied in massive or steeply dipping stratiform orebodies. For an inclined orebody, the inclination of the stope footwall must exceed the angle of repose of the fragmented rock in order to promote free flow of rock through the stope to the extraction horizon. Since stopes in these methods are unsupported, the strength of the orebody and surrounding rockmass must be sufficient to provide stable walls, faces, and crown for stope excavations. Additionally, the orebody boundary must be regular to minimize dilution. Due to the blast hole drilling and blasting technique, the minimum orebody width for open stoping is approximately 6 m (Brady and Brown, 2006). Pillar recovery is a common practice in open stoping, made possible by the use of backfill placed into primary stope voids. The backfill replaces the support provided by the ore pillar, allowing for pillar extraction.
Artificial support in mine openings is intended to control both local, stope wall behaviour and near-field displacements. There are two main categories of artificial support for ground control: mechanized support (e.g. rockbolts) and backfill. Potentially unstable rock near an excavation boundary may be reinforced with rockbolts. Backfill is used to fill stope voids and can prevent the progressive disintegration of near-field rockmasses in low stress conditions (Brady and Brown, 2006). Artificially supported methods include bench-and-fill stoping, cut-and-fill stoping, shrink stoping, and vertical crater retreat (VCR).
Cut-and-fill stoping is a very versatile method that can be adapted to an orebody with any shape (Bullock and Hustrulid, 2001). It is a very selective mining method that most commonly advances up-dip in an inclined orebody. Mining costs are relatively high compared to other methods; recovery is also high, and dilution is low. As such, it is an appropriate method for high grade orebodies (Bullock and Hustrulid, 2001). Cut-and-fill is a very controlled cycle of mining that is repeated many times in a single deposit. The simplified steps are:
- Drilling and blasting, where a 3 m thick slice of rock is stripped from the crown of the stope;
- Scaling and support, where loose rock is removed from the stope crown and walls and lightweight support is installed;
- Ore loading and transport, where ore is mechanically transported in the stope to an ore pass; and
- Backfilling, where a layer of backfill with a depth equal to the thickness of the ore slice is placed on the stope floor (Brady and Brown, 2006).
The success of this method depends on continued stability of the rockmass surrounding the work area where miners work continuously. This is achieved through controlled blasting, application of local rock support, and more general ground control using backfill. Cut-and-fill stoping is applied in veins, inclined tabular orebodies and massive deposits. When mining a large enough orebody, mining can be divided into multiple sections separated by vertical pillars. This method is suitable for orebodies dipping 35-90 degrees in either shallow or deep locations. The backfill allows for a weaker country rock, but the orebody itself must be a competent rockmass (Brady and Brown, 2006). However, if the orebody strength is very poor, a variation on cut-and-fill, underhand cut-and-fill, may be used (Bullock and Hustrulid, 2001). The ore grade must be sufficiently high to withstand dilution from backfill, but the grade can also be variable since lenses below the cut-off grade can be left unmined (Brady and Brown, 2006).
Case Study: Kristineberg Mine
Cut-and-fill mining is the primary mining method at the Kristineberg Mine, which is located in northern Sweden, approximately 130 km west of Skellefteå. The orebody is a typical vein structure with a dip varying between 45° and 80°. The host rock is a schistose sericitic quartzite and the rock immediately adjacent to the orebody is often highly altered, frequently very weak, talcy sericitic schists that vary in thickness between 0 and 3 m. The rock strength of both wall rocks and ore decreases from the hangingwall to the footwall as a result of metamorphic folding and faulting. Ground control problems, including roof collapse and wall slabbing, arose from a combination of variable rock quality and high in-situ stresses. Ore in the roof of the backfilled stope is subjected to large horizontal stresses, which results in the failure of both the roof and sidewall. In response to the resulting decline of ore production in the 1980s, extensive investigations of the feasibility of cut-and-fill mining at greater depth under these conditions were undertaken. The continuation of successful mining at greater depths is a result of the following findings:
- Dense support is necessary to maintain stability, and efficient, mechanized support is necessary to reduce costs and maintain reasonable production capacity; and
- The combined effect of changes in support strategy, more efficient support capability, and reduced level intervals has resulted in increased production reliability and capacity, with improved mining costs (Krauland et al., 2001).
Bench-and-fill stoping is a more productive alternative to cut-and-fill where geotechnical conditions permit. Here, initial drilling and excavation drives are mined along the length and width of the orebody. Mining advances by the sequential blasting of production rings into the advancing void and the ore is mucked remotely from the extraction horizon (Villaescusa, 1996). Following mining, the stopes are backfilled to provide support for the stope walls. An example of bench-and-fill stoping geometry is shown in Figure 4. This method may be used at several scales and variants. In some cases it has become the preferred method of narrow vein mining (Brady and Brown, 2006).
Shrink stoping involves vertical or subvertical advance of mining in a stope, where the fragmented ore provides both a working platform and temporary support for the stope walls, as shown in Figure 5. This method is similar to cut-and-fill stoping, where the fragmented ore fulfills a similar function to backfill used in cut-and-fill. It is generally applied to very narrow extraction blocks that have traditionally not been suitable for a high degree of mechanization (Bullock and Hustrulid, 2001). The suitable orebody type, orientation, geomechanical properties and setting for shrink stoping are virtually the same as those for cut-and-fill. However, the chemical properties of the ore become more important for shrink stoping where the rock must be completely chemically inert. The ore rock must also be competent and resistant to crushing during draw in order to maintain flow through the stope (Brady and Brown, 2006). Shrink stoping remains one of the few methods that can be practiced effectively with a minimum investment in machinery but is still not entirely dependent on manual labour (Hamrin, 2001).
Case Study: Mouska Gold Mine
Shrink stoping has been implemented at the Mouska gold mine, located 80 km west of Val-d’Or and 20 km east of Rouyn-Noranda. 72% of the ore is produced in shrink stopes, 20% from longhole mining, and the remaining 8% from development work of the mine infrastructure. The minimum width of the shrink stopes is 1.6 m, and three major joint sets are present in the rockmass. Most of the ground control problems can be attributed to (i) brittle failure of the diorite country rock under high stress; (ii) unstable blocks formed by the intersection of major joints; and (iii) major changes in orientation of the ore veins. The choice of shrink stoping at Mouska is based on the following factors:
- It is a selective method that allows daily assessment of the orientation of the vein being mined;
- It is a flexible method that permits better recovery of the ore in the extremities of the stopes;
- Ore inventory left in the stopes during the mining phase provides additional wall support; and
- Pillars can occasionally be left inside the stope, improving ground stability and mining grades (Marchand et al., 2001).
Vertical Crater Retreat (VCR) Stoping
Vertical Crater Retreat (VCR) stoping is a larger scale variant of shrink stoping, made possible by advancements in large-diameter blasthole drilling technology and explosive design. It is applicable in many places where shink stoping is feasible, except for orebodies that are narrow in width (less than approximately 3 m). It is particularly suitable for orebodies where sublevel development is impossible (Brady and Brown, 2006).
Unsupported mining methods include longwall mining and caving mining. These methods are distinguished from other mining methods because the near-field rock undergoes, by design, large displacements where mined voids become self-filling. Longwall mining is classically used in the deep mines of South Africa, where the near-field rock is usually strong and in-situ stresses are high (Brady and Brown, 2006). Several caving mining methods are generally well suited for massive ore bodies, including iron ore, low-grade copper, molybdenum deposits, other massive sulphide deposits, and diamond-bearing kimberlite pipes (Hamrin, 2001). These include block caving, sublevel open stoping and bighole stoping, and sublevel caving. Cave mining refers to all mining operations where the orebody caves naturally after undercutting and the fragmented material is recovered through drawpoints. Cave mining has the lowest cost for underground mining, provided the drawpoint size and handling facilities are appropriate for the caved material in a given mine (Laubscher, 1994).
Longwall mining is best suited for thin ore deposits that have a large horizontal extent. Ground support is used to maintain the excavation opening near the face, while the hangingwall behind the excavation can be allowed to subside. Hydraulic props, cribs, and pillars of timber or concrete are common ground support systems (Hamrin, 2001). This method can be used for both hard rock mining of metal ore and coal mining in soft rock. In both cases, the method maintains continuous behaviour in the far-field rock. An orebody must be dipping less than 20° and have a relatively uniform grade distribution. Additionally, any displacement along a fault must be less than the thickness of the orebody. In hard rock mining, longwall mining aims to maintain near-continuous behaviour of the near-field rockmass, which requires a strong and competent hangingwall and footwall rockmass. A generic layout of longwall mining in hard rock is shown in Figure 6. In cases where only a single pass of the orebody is mined, movement and closure of the hangingwall and footwall occur as mining advances. Once the hangingwall and footwall are in contact, the ground stresses at that location are invariant with further mining. This allows for the use of lighter rockmass support in the vicinity of mining activity (Brady and Brown, 2006).
Sublevel caving is a true caving technique that seeks to induce free displacement of the country rock overlying an orebody. Mining progresses downwards in an orebody where each sublevel is extracted as mining proceeds, as shown in Figure 7. Since gravitational flow of the fragmented ore rock controls the ultimate yield, development of the caving rockmass and the setup of the drill headings are the important aspects of the mining method. Generally, sublevel caving is suitable only for steeply dipping orebodies, with reasonably strong orebody rock enclosed by weaker overlying and wall rocks. The average grade must be high enough to sustain dilution to amounts that are perhaps greater than 20%. This method results in a significant disturbance of the ground surface, limiting its application to areas with suitable local topography and hydrology. Close control of draw is required to limit dilution of the ore stream. Geomechanics issues are prone to arising in production headings as a result of high concentration of field stresses in the lower abutment of the mining zone (Brady and Brown, 2006). This method has been most commonly applied to mining magnetic iron ores that can be easily and inexpensively separated from the waste rock (Bullock and Hustrulid, 2001).
In block caving, disintegration of the ore and country rock takes advantage of the natural fractures in the rockmasses, the stress distribution around the boundary of the cave domain, the limited strength of the rockmasses, and the tendency of the gravitational field to displace unstable blocks from the cave boundary. A schematic of block caving is shown in Figure 3. This method is distinct from all others discussed thus far as the primary fragmentation is achieved by natural mechanical processes. Block caving is a mass mining method, capable of high, sustained production rates at relatively low cost per tonne. It can only be applied to large orebodies where the height exceeds approximately 100 m. Productive caving in an orebody is prevented if the advancing cave boundary spontaneously stabilizes into, for example, an arched crown of blocks. Important geotechnical factors to consider when evaluating the caving potential of an orebody include the pre-mining state of stress, the frequency and surface condition of joints and other fractures in the rockmass, and the strength of the intact rock material. The most favourable rockmass structural condition for caving is one that contains at least two prominent subvertical joint sets, plus a subhorizontal set (Brady and Brown, 2006). It should be noted that in high stress fields, it has been observed that too rapid a draw can result in the creation of rockbursting conditions (Bullock and Hustrulid, 2001).
A summary of the geotechnical factors for each underground mining method, including the suitable orebody geometries, orebody grades, orebody and country rock strengths, and depths are shown in Table 1.
|Method Class||Method||Relative magnitude of displacements in country rock||Strain energy storage in near field rock||Suitable orebody geometry||Suitable orebody grade||Suitable orebody, country rock strength||Suitable depth|
|Pillar supported||Sublevel open stoping||..||..||..||..||..||..|
|Artificially supported||Shrink stoping||..||..||..||..||..||..|
|Artificially supported||VCR stoping||..||..||..||..||..||..|
- Barton, N.R., Lien, R. and Lunde, J. 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mech. 6(4), 189-239.
- Bieniawski, Z.T. 1989. Engineering rock mass classifications. New York: Wiley.
- Brady, B.H.G. and Brown, E.T. 2006. Rock Mechanics for underground mining, 3rd Ed. The Netherlands: Springer.
- Brown, E. T. 2003. Block Caving Geomechanics. Julius Kruttschnitt Mineral Research Centre: Brisbane.
- Bullock, R. and Hustrulid, W. 2001. Chapter 3: Planning the Underground Mine on the Basis of Mining Method. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 29-48. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.
- Hamrin, H. 2001. Chapter 1: Underground mining methods and applications. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 3–14. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.
- Herne, V. and McGuire, T. 2001. Chapter 13: Mississippi Potash, Inc.’s, underground operations. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 137-141. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.
- Hoek, E. and Brown, E.T. 1997. Practical estimates of rock mass strength. Int. J. Rock Mech. Min. Sci. 34:8,1165-8,1186.
- Krauland, N., Marklund, P.-I., and Board, M. 2001. Chapter 37: Rock support in cut-and-fill mining at the Kristineberg Mine. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 325-332. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.
- Laubscher, D.H. 1990. A Geomechanics Classification System for the Rating of Rock Mass in Mine design. Journal of the South African Institute of Mining & Metallurgy, vol. 90, no 10, pp. 257-273.
- Laubscher, D.H. 1994. Cave mining – the state of the art. The Journal of The South African Institute of Mining and Metallurgy, 94(10): 279-93.
- Marchand, R., Godin, P., and Doucet, C. 2001. Chapter 19: Shrinkage stoping at the Mouska Mine. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 189-194. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.
- Villaescusa, E. 1996. Excavation design for bench stoping at Mount Isa mine, Queensland, Australia. Trans. Instn Min. Metall., 105: A1–10.
Written by J.J. Day, Dept. of Geological Sciences and Geological Engineering, Queen's University