Strip Mining

CHAPTER 10.8

Strip Mining Gary Dyer and Ken Hill

INTRODUCTION

constrained by a mining property boundary, limiting surface feature, geological discontinuity, or economic limit with the maximum length of some strips being in the kilometers where conditions allow. A form of strip mining where the strips are aligned downdip instead of along the strike is applied in some instances but is rarer as economics generally favor strike mining. Each strip working area is denoted as a pit with a strip mine consisting of a number of pits. Strip mining encompasses a number of different mining strategies, each a unique combination of pit configuration, equipment selection, and operating methodology. This chapter begins by outlining a generic strip mining process. It then goes on to discuss alternative pit configurations, equipment selections, and operating methodologies for each of four mining strategies. The strategies have been developed in some detail with a major case study example used to provide comparison and contrast. Additional examples are referred to for variants of note. The chapter closes with some discussion on what drives the selection of one strip mining strategy over another as well as a brief commentary on future trends in strip mining.

For minerals that lie within economic reach by surface mining, and when the deposit’s specific geometry, either tabular or bedded, allows or dictates, an approach to extraction called open-cast, or strip mining as it is more commonly referred to, can be employed. Although a range of commodities such as phosphate, bauxite, tar sands, manganese, and even industrial materials from quarries have been recovered in this manner, the most common deposits worked by strip mining are coal deposits, and for this reason this chapter will refer predominantly to coal. The economic depth for strip mining is dependent on realizing a margin between its unit revenue value and the unit cost to recover, driven by depth and deposit complexity. The lesser-value bulk commodities will typically only be economic to shallower depths, whereas high-value coking coal, such as that found in Australia’s Bowen Basin, and where the stripping ratio allows, shows some strip mines currently operating at up to 150 m of depth with potential to progress to more than 300 m. Strip mining is a bulk earth-moving operation making use of large-scale mechanized equipment. An attractive geometry for strip mining is a tabular deposit with coal subcropping at economic depth and extending laterally, either flat or gently dipping, and constrained vertically within a number of seams. Deposit continuity relatively free of geological structure or intrusions is also attractive but not a prerequisite, although a heavily disturbed and/or intruded deposit and the resulting economies may limit the selection of equipment and operating method. Strip mining is characterized by its method of waste material, or overburden, movement, which is placed almost entirely in-pit. An initial cut is made on the coal subcrop, called the boxcut, and the overburden is placed on a natural surface updip of the subcrop line. The exposed coal is mined out and successive cuts, or strips, are taken to progress the mining downdip with the overburden from each strip placed inside the previous mining void. Individual strip geometry is typically from 30 to 100 m wide and to the economically recoverable basal coal seam. A strip will extend along the strike until it is

GENERIC STRIP MINING

The strip mining process from first workings to final closure can be described generically in six steps: 1. Clearing and topsoil removal 2. Fragmentation 3. Waste removal 4. Waste placement (including soil restoration and initial revegetation) 5. Coal mining 6. Mine restoration, maintenance, and eventual closure Figure 10.8-1 provides a stylized schematic to show these processes. Each of these process steps will be discussed briefly as a prelude to discussing the four strip-mining strategies. Some additional operational activities that support the mining process have been covered briefly due to their influence. These include rejects and tailings storage or disposal, river and creek diversions, and water management.

Gary Dyer, Manager Strategy, BHP Billiton Mitsubishi Alliance, Brisbane, Queensland, Australia Ken Hill, Managing Director, Xenith Consulting, Brisbane, Queensland, Australia

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Clearing and Topsoil Removal Prior to first workings, a mining area is typically covered with vegetation and topsoil. The vegetation must be cleared, and the topsoil must be recovered and stockpiled for later use in postmining rehabilitation. Vegetation is cleared by track dozers and pushed into piles and disposed of by burning, shredding for mulch, burying, or a variety of methods dependent on local customs and laws. An examination is made as to the quantity and quality of topsoil, and a target thickness to recover is identified. High-quality soils may need to be separately identified and stored. Topsoil is removed with earthmoving equipment, typically scrapers. However, where thickness allows, dozers assisting excavators to load trucks may also be an efficient operating method. The topsoil is stockpiled in the vicinity of the future mining operation so that it may be recovered for postmining rehabilitation. For a mature mine with ongoing clearing and topsoil removal occurring in parallel to final rehabilitation, it may be possible, and is certainly desirable, to place topsoil directly in its final position without stockpiling. It is important that topsoil stockpiles are not compacted by machinery or excessive thickness as biological activity, and thus their value, will be adversely affected by the exclusion of oxygen and moisture from the topsoil mass. Fragmentation Following topsoil removal, there is a relatively small thickness of weathered and/or unconsolidated material overlying much thicker units of competent overburden. However, where there is a substantial thickness of weak material, it may be possible, or even necessary, to progress directly to a wasteremoval process. This is the exception, and so fracturing, or fragmentation, of the overburden in advance of mining operations is required so that it may be handled safely and productively. Two main methods of achieving fragmentation are dozer ripping and drill-and-blast, with method selection dependent on the strength of the overburden, its thickness, and the volumetric demand of the mine.

Dozer Ripping

For relatively thin overburden of less than 4 m and where rock strength allows, a dozer-ripping process may be used for fragmentation. Use is made of the track dozer tines and the machine’s mass to break the rock into a manageable size. Operational effectiveness is highly sensitive to rock strength, with high-strength rock leading to poor productivity, as well as maintenance downtime, due to adverse wear on the machine. Drilling and Blasting

Drilling involves the creation of holes in the overburden within which explosives are placed. Drill holes are laid out in a regular pattern, either square or offset, the spacing of which is determined by the diameter of holes being drilled and the desired explosives density per volume of rock to be broken, or powder factor. Drill holes are drilled to either the roof of the target coal seam or to an operational depth to match the waste removal operating methodology. Where the valuable material (i.e., coal) is variable in depth, the driller may drill into the coal to ensure correct hole depth and then put a short length of stemming into the bottom of the hole to protect the relatively soft coal from damage when the overburden is blasted. Otherwise, holes are measured to ensure that the correct depth has been reached, with wet holes being marked and pumped out if the depth of water warrants it. Poor depth control will lead to uneven pit floors and inefficient waste removal. A presplit may also be used where a line of holes at a tighter spacing is drilled along the line of the next strip’s highwall and is blasted separately from the main pattern. The intention is to deliver a smooth wall that is productive to excavate the waste back to and safe for personnel and equipment to work under during coal mining. A blast design is prepared based on the overburden strength characteristics, the wetness of the holes, and the desired direction of movement of the blasted material controlled by placement and duration of hole delays. Bulk explosives are placed in the drill holes by mechanical means with ammonium nitrate and fuel oil (ANFO) being the most

Strip Mining

common explosive used. Specialist explosives may be used for stronger overburden with water-resistant emulsions used in wet applications. As each hole is loaded with explosives, an explosive booster tied to a detonating cord is placed within the explosives near the bottom of the hole. The top 5 m or so of the hole is filled with stemming to prevent the blast energy from escaping at the top of the hole and instead direct the blast energy into the rock mass to fracture it (and in some applications to actually move it sideways). The cuttings at the top of the drill hole are often used as stemming material, but some mines import size-graded rock to deliver a superior blast-containment mechanism. After the area to be blasted has been loaded with explosives, the shot firer will connect each hole with detonating cord and time delays. A lead-in detonating cord is run out to a position of safety and, with guards in place and the area confirmed as safe to blast, the shot firer will initiate the blast. Electrically initiated detonators with millisecond delays may be preferred where sensitive vibration control is required. A good blast outcome is typified by an absence of flyrock, no excessive sound and/or vibration, no misfired holes, and good visible fragmentation with the rock mass traveling in the desired direction. Strip mining is unique in that the drill-and-blast process itself can be employed as an overburden removal process. As the overburden is to be placed into the mined-out void immediately adjacent, certain pit configurations and operating methodologies lend themselves to cast blasting. Cast blasting is where a powder factor and delay design is selected to purposely cause the fractured rock mass to heave in the direction of the mined-out void with large quantities of overburden, up to 30%, resting in final position. It therefore requires no further handling by mining equipment. This is a particularly economical method of overburden removal. Waste Removal After blasting, the overburden may then be removed. Each end of the strip is called an endwall. The wall of the strip that is to progress successively downdip and is cut in in-situ material, or prime, is called the highwall. The in-pit overburden, or spoil side, wall is called the low wall because of its flatter slope compared to the highwall. (Refer to Figure 10.8-1.) Cast–Doze Excavating

At shallow depths of <30 m, a combination of cast blasting and pushing with track dozers will provide the most costeffective method of waste removal. Cast blasting can place up to 30% or more of the prime overburden into a final position. Track dozers can then push the remaining overburden into final position. If dozer pushes become excessively long or uphill, and therefore less productive, it may be more economical to supplement the operation with an excavator and a shorthaul truck operation. The highwall will typically be cut to an angle of 65° with excavator assistance in cutting the batters back to hard material. The low wall will typically be placed at between 37° (the angle of natural repose for fragmented rock) and 45°. Effective rock fragmentation is critical to the success of this method, as blocky oversize overburden has a significant negative impact on the productivity of dozer operations. Poorly cemented overburden may be moved by scrapers. Dragline Stripping

Where scale and deposit geometries allow (up to 85 m or so in overburden depth, depending on the class of machine being

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applied), a dragline can provide a more cost-effective method of waste removal. This requires sufficient annual demand for overburden stripping and mine life (in excess of 10 years) to pay back a high-capital, low-operating-cost piece of equipment. Simplistically, the dragline operates by moving overburden, or prime, from the current strip and placing it in the previous mined-out strip as spoil. Figure 10.8-2 shows a diagram of a dragline identifying its main physical features. The dragline needs to be placed at a level from which it can reach the top of the deposit to be exposed. This will either be its dig depth or operating level if less than the machine’s maximum dig depth, although the dragline can handle overburden above its operating level via overhand methods. The total thickness of overburden to be allocated to the dragline therefore may consist of a number of passes made up of a combination of overhand and underhand operations and perhaps even repeated for multiple seams. Figure 10.8-3 shows a typical walking dragline operation. Highwall angles range from 45° to 65° and depend on the competency of the material to be excavated, as well as joint orientations. Offset benches in the highwall may be established for reasons of overall wall safety in terms of stability, to reduce risk to subsequent mining operations below the wall, or for operational reasons relating to the maximum reach of the overburden drills. Low-wall angles are typically cut at 45° up to the dragline operating level and then lay back at a 37° angle of repose to the peak of the dragline spoil. The dragline will advance along the strip, moving by dragging itself on a revolving tub by 1–2-m “steps” delivered by two large feet on an eccentric cam, in regular blocks, typically 30 m in length, but they can be longer or shorter depending on the dragline’s operating level and thus its ability to reach the toe of its excavated block. For a narrow strip width, the dragline is often placed directly into a final position, whereas for a wider strip, as is more commonly the case, the dragline builds temporary operating positions out of overburden from which it places waste from the previous block into final position. This is discussed in more detail later. Any overburden that the dragline has to handle a second time is termed rehandle. An efficient dragline operation will minimize rehandle, which is achieved by setting the dragline’s operating level as low as possible. Unfortunately, the lower the dragline operating level, the lower its direct overall spoiling volume capacity,

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Figure 10.8-3 Typical dragline operation

which will then have to be compensated for, either by employing spoil side rehandle passes or removing material in advance by truck methods. Both of these are expensive, so rehandle is often an outcome of balancing desired coal exposure rates with installed dragline capacity and overburden removal economics. Dragline operating geometries provide a relatively narrow operating envelope. For the larger class of draglines, their operating parameters are a dig depth of 65 m, a dump height of 55 m, and a dump radius of 95 m with an effective thickness of overburden allocated to the dragline system of around 85 m (Bucyrus 2008). Overburden outside of this envelope will need to be rehandled if the dragline is to deal with it; if it needs to be rehandled twice, it will generally be cheaper to remove by truck methods. Errors in design and unexpected geological impacts such as faults or discovered areas of geotechnical weakness may result in overburden being locally over-allocated to the dragline somewhere along the strip. This may result in the dragline becoming spoil-bound where it can no longer place all of the overburden it has been allocated to a final position using its planned operating methodology. This needs to be rectified either by ramping its bench higher to provide additional spoil room, or, if spoil fit is already maximized, by walking the dragline into the spoil and rehandling overburden even further away, via an elevated bench or pullback operation, to create room for the current strip’s waste. This is an expensive exercise and can cause significant unplanned delays in coal exposure rates and so is to be avoided. Mobile, track-mounted draglines are still popular in variable terrain where overburden depths do not exceed the digging capacity of these smaller machines. The annual operating capacity of a dragline is a product of its rostered time, mechanical and electrical availability, operating use, and productivity. As they are capital intensive (in excess of US$200 million per unit), draglines are typically rostered to operate 24 hours per day, 7 days per week. Their relative simplicity as a predominantly electrical machine leads to high availabilities, typically running in

excess of 90%. Apart from geotechnical failures, excessive rainfall, or mine-planning failures (lack of blasted inventory), operating use is also high, running typically in excess of 90% of available time. Productivity will vary depending on the operation, with large swing angles from dig to dump, high spoiling, poor fragmentation, and rework of the excavated face or bench resulting in lower productivities. Total volumes moved per annum will range from 15 Mbm3/ yr (million bank cubic meters per year) for the most common smaller-sized machines to >30 Mbm3/yr for the larger machines currently deployed. A dragline, depending on its size, will consume the equivalent electricity per annum to supply a small town of 3,000 to 5,000 people. Concentrating so much productive capacity in a single machine and in a pit configuration that requires vertical stripping, however, is a significant risk exposure, and there have been instances of catastrophic failure via machines “falling” into the pit because of bench failures, basic mechanical failure of key structural components (materials failure or operator error), or flooding. For mines relying on a single or a small number of draglines, this results in a significant business impact that is not quickly or inexpensively recovered from. Small efficiencies in dragline use lead to large cost savings, so most large draglines will be fitted with operating monitors. These assist the machine operator to select optimum swing angles, bucket-fill factors, and casting radii. Dozer-Assisted Dragline Stripping

The dragline can be supplemented by track dozers pushing prime material into the void where, depending on strip width, it can be rehandled to its final position by the dragline. This operation is called production dozing and has the effect of increasing the effective prime waste movement rate, and thus coal exposure rate, of the dragline system, but at the expense of increasing dragline rehandles. Nonetheless, where there is sufficient spoil side room due to the dragline being underallocated waste in terms of spoil fit for coal exposure rate reasons, it is a low-cost, low-risk way to increase the total thickness of overburden allocated to the dragline system.

Strip Mining

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Truck and Loader Stripping

For more-restrictive deposit geometries, shorter mine lives, and where a more-variable scale of operation is planned, the superior flexibility of truck-based waste removal methods may be suitable. Truck methods involve the use of trucks and loading equipment to dig overburden and dump it. Loading equipment is selected to match the trucks, whereas the truck and loader package is selected to match the task. Highwall batter angles are 60° for each bench in competent material, with flatter angles for poorer material, but often with 5-m berms between benches leading to much flatter overall highwall angles. Large bulk thicknesses of overburden, typically >15 m, will be removed by electric rope shovel and ultra-class (>300-t payload) trucks as the most cost-effective operation, provided there is a sufficient annual requirement for waste stripping and mine life (>10 years) to pay back the higher-capital, loweroperating-cost position offered by larger electric shovels compared to smaller hydraulic excavators. For thicknesses >20 m and up to 25 m, overburden can be pushed down with dozers to create a safe working face height. Beyond 25 m of total thickness, multiple benches are often created. These might be separately blasted or through-blasted depending on material competency to allow them to be run over with the trucks. Where ready supplies of electric power are not available or at contract operations, this overburden task will alternatively be handled by large hydraulic tracked excavators, either in shovel or backhoe configuration. Electric rope shovels work most productively on a relatively flat bench, so where the dip of the coal seams exceeds the effective operating grade for the electric shovel, a secondary wedge operation using an excavator and trucks may be needed. Truck-shovel operations can be used in an advancedbench mode on upper layers of overburden to prepare a working bench for a large dragline. The truck-shovel operation removes variable topography, leaving a horizontal bench for the dragline. The dragline then is able to operate uniformly and efficiently at its optimum digging depth. Meanwhile, the truck haulage routes are arranged so that a natural-looking final topography is created without the need for spoil rehandle. For overburden thicknesses <15 m, it will usually be more cost-effective to run with smaller hydraulic tracked excavators loading appropriately sized trucks. As overburden thicknesses reduce, smaller excavators may be selected. Below 2 m of thickness, it may become more effective to have a dozer ripping and pushing overburden up to the excavator, whereas sometimes a pure dozer method is employed where a void is available within a short push distance. Wheel loaders are rarely used to load waste, often due to their inferior economics in overburden operations. A typical truck and loader fleet will consist of a loading tool, sufficient trucks to match the total cycle time (load, haul, dump, and return) to keep the loading tool running continuously, and various earthworks support gear including dozers, graders, and water carts. When used to remove overburden above the dragline, the truck operation is often termed a prestrip operation. When used to remove waste between coal seams below the dragline operation, the truck operation will often be termed an interburden operation, whereas for thin waste bands it may be called a parting operation. Where the truck operation uncovers coal in its own right, either above a

Courtesy of BHP Billiton.

Figure 10.8-4 Electric rope shovel and truck operation

dragline or in a nondragline mine, it is more commonly termed a truck and shovel operation. The width of strip for the truck and loader operation will typically be set to match the dragline strips so that a regular release of dragline strips occurs. This can limit the effectiveness of the loading operation in narrower strips by reducing the operating room, increasing the influence of edge effects, and limiting the opportunity to deploy doublesided loading of the trucks, thereby reducing productivity. Wider strips, however, result in longer truck cycle times and increased work-in-progress in the pit that manifests as a larger invested working capital. Scheduling bottlenecks are also introduced by larger batch sizes via wider truck and loader strips. These effects need to be carefully evaluated depending on local conditions. Figure 10.8-4 shows a typical truck and electric rope shovel operation. The annual operating capacity of an electric rope shovel is a product of its rostered time, mechanical and electrical availability, operating use, and productivity. As they are capital intensive (in excess of US$20 million per unit), electric rope shovels are typically rostered to operate 24 hours per day, 7 days per week. Their relative simplicity as a predominantly electrical machine leads to high availabilities, typically running in excess of 90%. Apart from geotechnical failures, excessive rainfall, mine planning failures (lack of blasted inventory), or sufficient trucks in the cycle to match the haul duty required, operating use is also high, running typically in excess of 90% of available time. Productivity will generally vary less than a dragline because of the relatively generic nature of electric rope shovel operation most of the time. Productivity will be affected when digging less than optimal bench height, ramping in or out of an area, or when poor fragmentation and difficult materials such as clays result in carry back within the bucket leading to increased truck loading times. Total volumes for electric rope shovels typically range from 15 Mbm3/yr to 20 Mbm3/yr for the larger machines currently deployed with the variability driven by the specific application. Three electric rope shovels will consume approximately the same electricity per annum as a large dragline. Electric rope shovels are inherently flexible but can present some constraints in terms of minimum required working areas, maximum operating cross-grade, and the requirement to manage a cable interacting with blasting operations and a fleet of trucks.

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Continuous System Stripping

For pits that require large annual quantities of waste movement (>15 Mm3/yr) to be placed via long haul cycles (>30‑minute return), it may be cost-effective to deploy a continuous or conveyor waste system, either by itself or in conjunction with other waste removal methods. Relatively few of these operate globally, and their successful implementation is highly dependent on developing a specific and robust mine plan that marries overall pit configuration, equipment selection, and operating methodology to the fixed nature of these systems compared to other mobile overburden-moving equipment options. These systems comprise a crusher station at the dig end, a long conveyor (>5 km), and a spreader at the dump end. The crusher may be fed by trucks or directly by track dozer, excavator, shovel, bucket-wheel, or even dragline. These systems are capital intensive, being comparable in capital cost to a dragline with similar annual prime capacity but with an operating cost that sits between a dragline and truck and loader systems. For these reasons, only the longest-equivalent truck hauls provide an effective cost offset; economic payback can take a number of years and depends on achieving large annual volumes, which means minimizing the magnitude and frequency of partial- or full-system relocations. This is a developing mining strategy that will see more potential for economic application as pits deepen, but it is also under competition from emerging technology step-change improvements in truck and loader systems, such as partial or full automation. The environmental advantage of continuous system stripping is the ability to restore spoil in an approximation of the original strata profile. Waste Placement The movement of waste is a pure cost to the mining operation and has no direct economic benefit, so the placement of waste will generally be driven by a least-effort approach. For equipment other than trucks and conveyors, this dictates the overburden only be moved a short distance within the practical and economical operating envelope of the equipment in question. This will typically be <200 m in the horizontal from the point of origin with changes in elevation generally <50 m. Cast–doze excavation, dragline stripping, and dozerassisted dragline stripping methods generally place the overburden into the void from the previous strip either directly or with some rehandle. Minor truck stripping operations engaged in either parting or thin interburden operations, and generally at or near the bottom of the strip, will seek to minimize the haul. Overburden is thus placed inside the strip being mined or nearby, perhaps being used to regrade an access ramp that may be running at a flatter grade than that required by the coal mining trucks. Bulk truck stripping operations, not supporting a dragline stripping method, also place the overburden in the previous void. Least-cost operation is generally achieved when material is hauled on grade, or as flat as possible, and it is desirable to establish the dumping strategy to sustain an average haul. This will ensure that short hauls do not idle spare trucks unduly and long hauls do not create instances where the loading equipment is undertrucked and production is lost. It is vital to have an established road system that allows access to multiple dump levels, and this road system will grow in complexity with pit depth and thus the number of dump lifts being developed. For bulk truck stripping operations supporting a dragline stripping method, otherwise known as prestrip, the overburden

is placed on top of the dragline spoil. The proximity of the truck dump to the active spoil area is determined by the geotechnical competence of the waste material itself and the strip floor; however, a good general rule is two spoil peaks, or prior strips, back. Dump geometry is dictated again by geotechnical competence but also the location of roads to access the dump and the desired intensity of dumping operations that need to be designed for. A typical overall operating angle of 19° for the active dump face is typical for truck dumping, and this profile is the result of a series of dump lifts that batter down at angle of repose interrupted by catch berms and roads. The economic cost of elevating overburden with trucks relative to horizontal haul is approximately 20:1, which is to say a dump will preferentially develop horizontally until economics dictate that an additional dump lift be added. For this reason, the dumps for mature pits will adopt a typical profile with the greatest height in proximity to the active dig area with a trailing back of the dump. At this point it is worth considering the geometric drivers of dump development as well as economics. Figure 10.8‑5 shows how the dumping strategy changes over time as a strip mine deepens. As the pit progresses downdip, batter effects demand greater and greater volumes of overburden to be moved, and this generally reports directly to the truck operation as the dragline is allocated a fixed thickness of overburden for each strip. On the dump side, the dump end batters play a similar role but have the inverse effect: as the dump gets higher, it has a correspondingly lower volumetric capacity due to less pit length available for dumping. The system dynamic is simplistically one of an inverted cone-shaped pit getting deeper, filling a conical stockpile that has a finite capacity for a fixed basal area. After the dump reaches capacity, only three responses are available. First, any perpendicular ramps to the base of the strips that create large valleys reducing available dump room can be filled with alternative clearance of the mineral to be mined to occur by parallel low-wall ramps or via a highwall ramp system. Second, the dump can be extended in the direction of the original commencement of mining. This can have grave consequences for any progressive rehabilitation that was undertaken without a final landform design and can lead to redisturbance of previously rehabilitated areas. The same issue might also apply to any infrastructure that was also placed without due consideration of long-term dump designs (e.g., coal haul roads, power lines, stockpile areas). Third, overburden can be placed out of pit or, in other words, beyond the limits of the excavated-to-date pit shell. In some regions, this out-of-pit material is called excess spoil. This can be expensive, involving long hauls or placing overburden on the highwall side of the pit and on top of future resource areas. This may be undesirable unless there are sterile areas due to geological disturbance, intrusions, or underground operations or other factors that may preclude strip mining in that area, or unless the subsequent rehandle of this waste at a future date is still economical. An additional downside to out-of-pit dumping is that it increases the total area of mining disturbance and creates an additional net area to be rehabilitated and maintained after mining ceases. Excess-spoil dumps have created environmental challenges whenever they intrude into established natural drainage channels. The placement of overburden for deep strip mines is an emerging problem and generally starts to manifest at depths of around 150 m. A further uncertainty and risk is the geotechnical stability of large in-pit dumps and is the subject of ongoing research.

Strip Mining

(A)

(B)

8-km-long pit with a ramp every 1 km. At Strip 1 the pit is 40 m deep and the dump is 20 m above natural surface.

At Strip 30 the pit is 130 m deep and the dump is 105 m above natural surface. The ramps have been partially filled and regraded to 10% slope.

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At Strip 55 the pit is 210 m deep and the dump is 125 m above natural surface. Every second ramp is now closed and filled above original ground level, and active ramps are now 2 km apart.

At Strip 70 the pit is 270 m deep and the dump is 155 m above natural surface. Every second ramp is now fully filled to the top of the dump.

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Figure 10.8-5 Dumping strategy development with depth

For an integrated truck and dragline method, an additional operational complexity to be resolved is the interaction between the trucks and the dragline with both equipment types often having to operate in close proximity at the same time. The truck operation will seek to minimize its average cycle time by crossing the open strip when it is excavating near the middle of the strip to get to the dump, but the dragline operation seeks to operate long continuous strips for maximum efficiency. One compromise is the use of cross-pit bridges that the trucks travel across. Temporary bridges are constructed from blasted overburden and built to the height of the dragline operating level. When the dragline is a suitable distance away from the temporary bridge, the trucks use it to provide a short-haul access across the pit. They are “rolled” by the dragline on each strip to recover the coal that lies beneath them. The downside of temporary bridges is that the bridge is not always available for use by the trucks, or its use by the trucks becomes a constraint on the mine schedule, and the rolling of bridges has a productivity impact on the dragline. An alternative approach is the use of permanent bridges. These are constructed from

in-situ material that is left in place in each successive strip. The coal beneath them is effectively sterilized as the successive mined-out strips are filled with overburden that either buttresses or buries the remnant mineral material. The bridges may be built to any height and may incorporate multiple crosspit roadways at different levels to further optimize trucking operations. The downsides of permanent bridges are the sterilization of resources, which is magnified for higher bridges due to larger bases from the batter effects, as well as their role in limiting the length of strip the dragline operates with the corresponding productivity impact from end effects and more frequent machine relocations. The rest of the trucked overburden either travels around the end of the pit or up a highwall ramp system to a natural surface and then around to the dump. A major advantage for endwall road systems is the provision of cost-effective on-grade highwalls, with a detraction being additional stripping demand from access roads cut into sterile ground, which can be substantial at depth, especially for short strips. A major advantage for highwall ramp systems is that they

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are excavated in material that is part of the ongoing mining operation, and generally already exist if coal seams are present within the prestrip working horizon. A detraction is that productivity impacts the excavation operation from removing and re‑creating ramps (if not already in place) as well as the potential need for loaded trucks to haul up to a natural surface and then haul downhill on the dump side, which is inefficient. Placement of overburden by conveyor systems is most similar to truck operations. A spreader constructs the overburden dump. Depending on the equipment configuration, the overburden dump may be built in benches radially or along the strike. An upper stacked overburden bench of 15-m height may be built simultaneously with a lower filled overburden bench of up to 45-m thick before the whole system is advanced. Because of the immobile nature of continuous waste systems, the conveyor will either travel around the end of the strip or across a central permanent bridge. Figure 10.8-6 shows a diagram of a typical spreader dump. Landform Strategy—All Placement Methods

The waste dump, or landform, strategy is dependent on a number of key factors including the long-term swell factor, operating pit and progressive and final landform geometries, any mining or dumping constraints, the total static volume balance, and the dynamics of optimal dump construction. The influence of the final pit limit and final pit treatment should also not be underestimated, with impacts for deep pits likely to affect the entire postmining landform constructed early in the mine life, but not manifested until the latter years of the mine life. Variations in the elevation of spoil dumps due to swell factor and the inherent displacement by draglines of contour highs and lows may lead to substantial drainage mismatch between disturbed and undisturbed lands. Surface contouring software can generate plans for final restoration that respect geomorphological principles. Coal Mining Although the overburden removal process accounts for the bulk of the operating expenditure for a strip mine, the coal itself is the reason for the stripping and provides the revenue to support overburden removal. The overburden and coal seams are physically coincident at their boundary, so it is inevitable that imperfect mining

processes will result in some unplanned removal of coal with the overburden, or loss, and some unplanned addition of overburden to the coal, or dilution. Typical losses and dilutions by mass can range from 5% to 10% with higher losses incurred for thinner seams and dependent on local practices and conditions. Loss is directly incurred through blasting that can cast some of the coal into the final spoil position to be lost forever, but also blast-enabled through the disturbance of the overburden–coal interface. By using large-scale earthmoving equipment, driven by the paradigm and key performance indicator (KPI) of minimizing unit operating cost, removing the overburden near the coal can easily be biased toward coal loss. An effective remedy can be to set coal recovery as a KPI or deploy a more selective equipment configuration when approaching the overburden–coal interface in the last 1–2 m. This incurs a higher unit cost for overburden removal, but the affected volume is quite small and the total cost impact minimal. Similarly, the coal mining operation is often biased toward loss over dilution due to the high visibility of overburden within the mined coal. For high-value coking coal where a processing plant is available to remove dilution and the cost of replacing lost coal is high, it will generally be economical to favor dilution over coal loss. For lower-value coal such as thermal coal, which generally has a lower cost to produce and is often crushed and sold without washing, it is usually more economical to favor coal loss over dilution. Some mines reduce dilution by use of a commercial road sweeper on the coal surface to windrow rocky material away from the coal-loading operation. Coal mining is commonly by loader and truck systems. For thin seams or soft coal it is dug unblasted or may be ripped and pushed by dozers or even graders as part of the mining process. High-value, thin-seam coals have been excavated by tarmac recovery machines called fine graders or special continuous surface miners whose operation combines the fragmentation and mining process. For thick and/or hard seams a square drill-and-blast pattern is employed to fragment the coal prior to mining. Loading equipment used is either a frontend wheel loader, a diesel excavator in shovel or backhoe configuration, or, more rarely, a smaller electric rope shovel for some mines that have particularly thick coal seams. The choice of loading equipment will be driven by the required mining rate and the thickness of the coal seam and its continuity, with excavators in backhoe configuration being the digging tool of choice for highly faulted or banded coal seams, where selective mining is required. Trucks will either be rear dumps, electric or diesel, or belly dumps. Rear dumps have the advantages of being more maneuverable and faster up the coal access ramps compared to belly dumps, as well as being able to handle ramp grades of up to 10%, whereas they have the disadvantage of being slower on the flat with top speeds of 65 km/h compared to 75 km/h for belly dumps. For very long flat hauls (>20 km), and where ramp grades allow, highwayadapted road trains consisting of multiple self-tipping trailers with up to 300-t capacity are used. Figure 10.8-7 illustrates a typical coal mining operation. Depending on blending requirements, coal may be directmined and fed to the wash plant continuously from a number of strips in different pits at once. Alternatively, it may be mined to stockpile for reclaiming later, either by a mechanical stacker reclaimer system or by manned equipment such as a batch-fed process.

Strip Mining

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Figure 10.8-7 Coal mining operation

Mine Restoration, Maintenance, and Eventual Closure The objective of mine closure is to return to the wider community a postmining or final landform that is safe, stable, and sustainable. Safety dictates that the final landform does not present a risk to humans or wildlife. Stability dictates that the final landform is both erosionally and geotechnically stable, with erosional stability more challenging to achieve. Sustainability dictates that the final landform reduces and prevents pollution, enhances biodiversity, and delivers a beneficial use with no ongoing liability to the company that owned and operated the mine. Strip mining by definition disturbs large areas of land, so the greatest impact, and ultimately closure cost and potential trailing liability, relates to the creation and maintenance of the final landform from the open voids consisting of final strips and ramps. Other aspects of mine closure include the demolition and rehabilitation of mine and fixed plant infrastructure, but these are small contributors compared to the landform. In some jurisdictions the final landform is highly prescribed, whereas in others, legislation provides broad guidance within a set of principles. Final landforms can range from voids whose batters have been regraded to voids that have been fully backfilled to the original topographic levels. In a strip mine it is worth noting that, by virtue of the operating methodology, mine closure in terms of the delivery of a final landform should be seen as a progressive activity that commences the day the mine opens and occurs more or less continuously until the site is fully rehabilitated. For this reason a strategic landform study should be undertaken as part of the initial development study, or soon after, so that the construction of progressive landforms can be undertaken in a cost-effective manner, maximizing the design of truck dumps to emulate as near as possible the final landform design. This will involve the bulk placement of material to progressively

deliver the final landform fill, preferential placement of adverse material types such as acid-forming or highly erosive spoils within the spoil mass, preferential storage with progressive use of materials suitable for controlling erosion of external surfaces, and topsoil to aid in revegetation of the regraded spoil mass to deliver the final landform finish. The other aspect of the final landform is the treatment of the final voids consisting of access ramps and the last mined strip. Significant costs and liabilities can be incurred if closure is not approached in a planned way. However, there are also significant operating benefits to be accessed from developing a comprehensive operating-to-close strategy and implementing it in the last 5 to 10 years of the mine’s life. For deep pits it is likely that any perpendicular coal access ramps will have long been closed for dump room reasons, but for shallower pits approaching closure, and where still in use, they can represent a low-cost dump opportunity. Final strip voids become more valuable as a low-cost dumping opportunity for adjacent pits than as a source of marginally economic coal. A mitigated closure strategy will generally see all ramp voids, and most final strip voids, filled with waste from the remaining operating pits. In time these pits themselves will self-fill as a way to reduce unit operating costs and still yield a positive cash margin, albeit with ever-reducing total mine output. Where sufficient area exists and economics allow, some mines have wheeled their pits around in a large set of 90° offsets so that the last pit is adjacent to the first pit and thereby fills it. Regulating authorities in some countries demand a final landform design and progressive delivery as a condition for operating a mine, whereas in other countries the approach is less prescriptive. Nonetheless, the final landform is an integral part of operating a strip mine, and it is critical that a strategic landform plan is developed early and followed throughout the mine life.

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Cost per Product Metric Ton

$100 $80 $60 $40 $20 $0

35-m Dragline

70-m Dragline

70-m Dragline + 50-m Truck & Shovel


A. Cost contribution by process

Total Cost Contribution

100% 80% 60% 40% 20% 0%

35-m Dragline

70-m Dragline



B. Cost proportion by process

Truck & Shovel (12 trucks)

Drill & Blast

Truck & Shovel (6 trucks)

Coal Mining, Processing, Train Loading General Overhead

Dragline

Figure 10.8-8 Relative cost by scenario

Mining companies are able to operate by virtue of the license to operate (LTO) they hold. A company’s LTO is a product of its adherence to statutory requirements and the literal permits this provides in combination with figurative permissions provided by meeting the expectations placed on it by the wider community. At its broadest definition, this community can consist of local, regional, national, and international stakeholder groups comprised of individuals, institutions, and organizations. The strength of an open-cast mining company’s LTO is directly related to how it manages progressive and final closure and specific mine planning and design aspects such as landform, treatment of final voids, water drainage and quality, and how they ultimately impact postmining land use. Relative Economics by Business Process and Changes with Depth Having discussed the key operational tasks that define the strip mining process, it is worthwhile contrasting the relative

economics of each, how these relativities change with depth, and the operational implications of these movements. Four scenarios were developed based on a single 10-m coal seam, increasing waste depths, and differential waste allocation by waste-stripping process. Costs include both capital and operating components. Figure 10.8-8 highlights the contribution waste removal makes to total mine gate cost for the four scenarios—dragline operating productively at shallow depth (35 m), dragline operating at its effective limit (70 m), 50 m of truck and shovel prestrip introduced with relatively short haul (6 trucks), and 100 m of truck and shovel prestrip introduced with relatively long haul (12 trucks). Some key points become clear. For shallow deposits (35 m), coal-related processes make up a significant proportion of total cost, and the operation will be oriented around minimizing coal mining cost. This suggests frequent coal access ramps with coal haul roads optimized for haulage to the processing plant with management focus on the coal mining process. As the seams progress deeper (70 m), drill-and-blast makes up a more significant proportion of cost, and focus will shift to optimizing drill-and-blast operating practices. Dragline operations are now the majority of total cost, and there will be extreme focus on productivity and equipment uptime. Beyond 70 m, as truck and shovel prestrip is introduced, it will work around the dragline process, and its role will be to ensure costeffective dragline operation. At extreme depths (100 m of truck and shovel, and 70 m of dragline), waste stripping dominates the cost structure with truck and shovel operations now contributing as much cost as all of the other operations together. The coal mining and dragline operations will work around the truck and shovel operation to ensure its cost-effectiveness. The mine design will be modified to optimize truck and shovel at the expense of all else. The scenario also shows the additional cost loading due to task—the additional depth has a linear impact providing that the same haul cycles can be maintained, but simple dump geometry dictates that with depth, not only is there more waste, but it also has to be hauled farther and higher. The additional cost from this increased task is shown in Figure 10.8-8 as the 6-truck to 12-truck cost increment, with the result being an exponential cost increase. At this point, attention will be given to seeking out optimized dumping practices, new dump locations (potentially on top of high-cost coal or future underground areas), and innovative waste movement methods (conveyors being one example). Consideration at this point should be given to the relative cost of underground operations as an alternative means of generating raw coal for processing. A 4-m-thick seam at around 70-m depth will be exposed by dragline for approximately the same operating cost per metric ton as a longwall operation and for around the same mining capital investment. The choice of approach around this approximate opencut stripping ratio of 12:1 prime to raw coal (9:1 for a pure truck and shovel operation) will be governed by practical considerations. Opencut mining is inherently lower risk and has a wider envelope of application given that underground operations can be severely impacted or even excluded by factors such as geological faulting, intrusions, weak immediate roof, seam gas, seam water, and so on. Other Operational Aspects Three other operational aspects are worth discussing in respect to their unique impact on strip mining. These are the storage or disposal of washery rejects and tailings, river and creek diversions, and water management.

Strip Mining

Rejects and Tailings Storage or Disposal

Typical product yields range from 50% to 90% by mass for all material fed into the coal washery. Coal washery by-products, the remaining 10% to 50% of the mined coal mass, include coarse reject and wet tailings with the bulk by volume being the coarse reject material. Coarse reject is often stored in close proximity to the wash plant, either in a stockpile on a natural surface or, if available, a completed pit void, and it is placed there either by conveyors or trucks. Tailings are either placed in a conventional wet tailings facility constructed as a dam, or codisposed with reject in either a dam or completed pit void. Wet tailings facilities have large footprints, are high in capital to construct, are expensive to rehabilitate, and carry significant environmental risks both during their operation and postrehabilitation. For this reason, a completed pit can often be seen as an ideal storage and disposal solution. Deposit geometries, however, generally do not offer up completed pits early in the mine life, as a strip mine will work the entire strike on a cost equivalence basis for the entire mine life. Little economic delineation exists between pits that are often arbitrary subdivisions of what is usually a single deposit broken up by operational needs such as coal access ramps or creek corridors. More commonly available are final ramp voids left behind as the deep operation has moved to fill the mouth of a ramp as a short-haul dump option. Unfortunately, a competing alternative use for any final voids, either ramps or completed pits, is a potential low-cost dump, and given the cost ratio between horizontal and vertical haul, it is surprising how far away a truck system can reach economically when its alternative is to elevate overburden. For this reason, the best economic use of voids may not be to fill them with wet tailings, especially when water itself for many mines is becoming a scarce and valuable commodity. In these cases, a better alternative may be a partially or fully dewatered tailings paste or cake mixed with reject and disposed directly into the dump mass where it is contained and the long-term liability is minimized. No burial of washery wastes into the reclaimed overburden should occur without a thorough analysis of the wastes’ potential for harm to the restored groundwater regime. This may be, among other things, an acid–base accounting or a column test for long-term solubilities. River and Creek Diversions

Most strip mines are tens of kilometers long with few reasons to break a continuous strip into separate pits other than arbitrary subdivisions based on the needs of the mining operation; infrastructure corridors for road, rail, power, or water pipelines; and so on. It is highly likely that in this long strike there will be a number of water courses traversing planned mining areas. It then becomes an engineering and strategic analysis to weigh up the positive value of the coal that lies below a creek or river and the dump space created by removing a valley through the future maximum dump envelope against the negative value from the initial capital cost of a diversion, the ongoing operating cost to maintain the diversion, a final capital cost to reinstate the water course in or near its original course if applicable, and an estimate of the risk of exposure to a trailing liability to maintain the diversion into perpetuity. The nature of a strip mine means that it is likely that any permanent diversion will be quite long, whereas a temporary diversion can just shift the problem to another pit with the commensurate risk of reestablishing a water course through

999

what is now effectively a postmining landform that is more susceptible to damage from water erosion. Water Management

Because of their long strike and progressive development over time, mature strip mines can carry large internal water catchments. Environmental licensing conditions relating mainly to salt levels and turbidity can often mean that poor-quality mine water is not easily discharged off-site. For this reason, mine water is stored and either evaporated or partially consumed in the water-based coal processing operation. The large volumes that can accumulate over time or from single intense rain events often mean that a sacrificial pit is selected as the only economically viable water storage solution. The pit may be used for a number of years until it becomes the most attractive pit remaining on-site, at which point another pit is selected for water storage and the first pit is pumped out. After the water is removed, there is often a thick layer of mud in the bottom of the previously dormant pit that also needs to be removed. Mud removal is a slow and expensive process, and if not fully completed before stripping is recommenced in the affected pit, can lead to ongoing spoil stability issues. Massive spoil failure in a deep pit is prohibitively expensive to remediate. It is conceivable that for a very deep pit subjected to a massive failure, it would not be economical to recover the pit because of the large capital cost involved; thus, significant resource sterilization could occur if extraction by underground means is not deemed viable.

ALTERNATIVE STRIP MINING STRATEGIES

The four most common strip mining strategies will be examined as follows: 1. Cast, doze, excavate operation 2. Single-seam dragline operation with prestrip 3. Multiple-seam dragline operation with prestrip 4. Truck and shovel operation A strip mining strategy may be typified by a unique combination of pit configuration, equipment selection, and operating methodology with the combination essentially a response to the deposit being worked. For this reason, a methodical approach making use of the following discussion framework will be applied to examine each of the four major strip mining strategies deployed in mines today: • A brief generic discussion of the mining strategy • A specific mine case study used to examine –– Pit configuration—mine location and layout and the deposit’s geological orientation, –– Equipment selection—the physical and volumetric scale of extraction, and –– Operating methodology—an operational description in terms of overburden movement and coal mining • Reference made to variants of note as applied at other mining operations Other variants within the presented strip mining strategies are deployed where the specific circumstances warrant it. The combination of final and interseam waste removal and coal mining in thin-seam mines may be achieved by the application of continuous surface miners or graders. Steepdip mining, as in Colombian thermal coal mines or the United States anthracite district, gives rise to operational variations of the truck and shovel strategy, whereas mountain-top removal

1000


in the United States’ Appalachia region is unique and can be considered to be a multiseam modified area mining method. Cast, Doze, Excavate Operation “Cast, doze, excavate” refers to a strip mining technique whereby the overburden is cast blast, followed by significant removal by large bulldozers, and the remainder of the overburden is removed using an excavator and trucks, or sometimes by draglines. This method is most applicable to relatively shallow deposits. In these operations, the overburden blast is designed to not only fragment the overburden to enable removal, but also for the blast to significantly move the overburden into the adjacent open void, reducing the amount to be removed with mining equipment. This style of overburden blasting is referred to as cast blasting. Because of the increased blasting vibrations and noise, the method should not be used near dwellings and residential districts without first determining potential dust, noise, and vibration levels. Bulldozers are more productive and hence more economically attractive for combinations of relatively short push distances and fairly flat grades. Where bulldozers can be used in this manner they are usually lower total cost than truck and loader systems. In an ideal situation, the bulldozers are used to remove the overburden until the push distance and grade combination is no longer lower cost than the next alternative (either truck and loader systems or draglines). In some situations, cast doze is used to remove all of the overburden material. Case Study: Groote Eylandt Mine

Groote Eylandt is located off the coast of east Arnhem Land in the Gulf of Carpentaria about 640 km from Darwin, Australia. The mining leases are located on the western plains and cover 84.5 km2 or 3.74% of the island’s area. Pit configuration—Geological orientation. The Groote Eylandt ore body occurs as a subhorizontal sedimentary layer, gently undulating and dipping to the west. It is a continuous horizon, ranging in thickness from a few centimeters to 9 m, consisting of hard, high-grade, cemented pisolitic and massive manganese oxides. The overburden ranges in thickness from 0.5 to 35 m and consists of lateritic overlying clays and gravels. The ore body is vertically zoned and can be mined as two distinct layers: 1. A middle mining horizon consisting of massive mangite, cemented mangan-pisolite, or loose mangan-pisolite 2. A bottom mining horizon consisting of siliceous mangite The beneficiation plant is located in the middle of the mining leases, with 90% of the resource within 8 km of the processing plant. Equipment selection—Physical and volumetric scale of extraction. The mine typically removes 18 Mbm3 of overburden each year, comprising 0.4 Mbm3 of topsoil, 17 Mbm3 of dozer production, and 0.6 Mbm3 truck and excavator operation. This total overburden removal exposes 6.9 million t of run-of-mine (ROM) ore to be delivered to the primary crushing station to yield 3.6 Mt of manganese product. Current mining equipment is a mix of stripping dozers, excavators, and rear-dump trucks. The mining blocks or cuts are 40-m wide and 200-m long, with strips arranged to follow ore-body contours and to control groundwater.

Operating methodology. The Groote Eylandt Mining Company Pty Ltd. (Gemco) mine is a conventional, shallow, opencut strip mining operation involving the removal of manganese ore. The mining sequence is a continuous cycle; areas from which ore has been extracted are backfilled with overburden from the next strip to be mined before then being rehabilitated. This method results in the mining site moving across the ore body disturbing only a small section of land surface at any given time. To meet Gemco’s customer product-quality requirements, ore can be mined from a variety of pit sources. The first step is to clear trees on the planned mining blocks, using a bulldozer with a tree rake. The topsoil is then removed and returned to prepared backfill areas or stockpiled for future rehabilitation. Overburden is primarily removed using a fleet of Global Positioning System–equipped track dozers that push the material into the adjacent mined-out cut (strip mining technique). The dozers are used to remove all overburden, exposing the ore. Typically, the dozers work in two separate teams, each team working a strike length of 250 m to 500 m and a pit width of 40 m. At times they are used as one fleet to provide peak capacity for short-term ore requirements. Historically, scrapers were also used to remove topsoil and overburden when the overburden was significantly thinner. To reduce costs and handle a wider range of material types, production dozing was introduced. A small amount of overburden is removed using a truck and excavator method, and this method is applied in thicker overburden areas and where digging from above with excavators is used to better handle wet conditions. A rotary percussion rig is used to drill the ore horizon, and blasting is conducted using emulsions for both wet and dry holes. The blasted ore is removed using a hydraulic excavator. The excavator is also used for overburden stripping for initial or box cuts. Ore is trucked from the pits to the primary crushing stockpiles in 85-t-capacity rear-dump trucks. Some direct haulage also goes to the primary crusher, provided the ore type coincides with the ore treatment campaign through the plant. Approximately 22 stockpiles are maintained near the primary crusher and categorized according to grade and lump-tofines ratio. The haulage distance from the pit to the stockpiles varies from 2 to 10 km. In total there are about 30 km of haulage roads that are serviced by ancillary equipment. Rehabilitation of mined areas started in 1970. Seeds from approximately 25 native forest species are collected from within the mine lease. This seed is stored until the wet season when it is broadcast over areas that have been backfilled and covered with topsoil. Backfill areas are designed to maintain the general premining slopes. Single-Seam Dragline Operation with Prestrip For single coal seams that occur deeper than 30 m and where economies of scale allow, a single-seam dragline method is often applied. Although multiple seams may be recovered, for the purposes of terminology where only one seam is directly exposed by dragline operations, it is called a single-seam dragline operation. Because of their specific geometric limitations, efficient dragline operations are dictated by the interaction of the particular deposit geometries and the machine’s key working dimensions. A dragline moves overburden by a simple repetitive cycle involving filling the bucket through a combination of pulling in the drag rope and letting out the hoist rope, raising the bucket by the reverse rope movements,

Strip Mining

swinging to face the spoiling area, emptying the bucket by releasing the tension on the drag rope, and swinging back to return the bucket to the dig face or bank. Strip width is typically 50–70 m wide with the strip being placed into the adjacent mining void in discrete excavation blocks of 25–30 m in length, depending on the dragline operating level. A number of different dragline methods are available to place the overburden in the previous void. Selection of method is driven by a number of factors: • Deposit geology—The number of seams to be uncovered by the dragline and the overburden and interburden thicknesses • Geotechnical stability—Certain materials in the strata may be either unsuitable to place the machine on or unsuitable to place in certain locations within the spoil • Spoil fit—The machine will need to sit low enough to reach the top of the coal, but it will also need to sit high enough to ensure it can place the spoil into its final position • Productivity—Will be maximized with lower average swing angles, but swing angle may be less of an issue if there are a large number of hoist-dependent swings (where the swing speed is less than maximum due to the machine waiting on the bucket to be hoisted to height) • Scheduling—At times it may be necessary to adopt a dragline method that may be less productive overall due to increased relocation times but that releases coal sooner and more consistently Dragline methods are analyzed in two- and threedimensional (2-D and 3-D) volumetric models as well as simulations. Two-dimensional analyses called range diagrams are used to check for theoretical spoil fit and calculate sectional rehandles. Figure 10.8-9 shows an example of a range diagram sequence. The section is perpendicular to the strip and shows where material is dug from (A), dumped to (A'), and the operational sequence (steps and fill type). Where topography is variable or the influence of certain design features are such that a 2-D analysis is unsuitable, a 3-D check for spoil fit and rehandle rates can be undertaken. In addition to the simple dig-to-dump balance, more sophisticated software models can be used to simulate the dragline machine itself to yield theoretical production rates in cubic meters per hour based on a dynamic analysis to determine the average swing angle and relative frequency of hoist and swing dependent cycles. A brief overview of the main dragline methods are given in the following paragraphs. A completed range diagram is shown in Figure 10.8-10. In the side cast method, the overburden is dug from in front of the machine to expose coal and dumped directly in the void from the previous strip (Figure 10.8-10A). It is most often used for shallow overburden thicknesses where the dip is relatively flat and there is no requirement for selective placement of overburden in the spoil. It is also used for multiseam operations for exposing the top coal seam as there is a large void to spoil the overburden into. The key bridge method (Figure 10.8-10B) is suitable for deeper deposits up to the dragline’s ideal operating envelope. Sequentially it involves excavating a block of overburden 30-m long to create the new highwall (the key) and using this overburden to build a bench out into the previous strip’s void. The dragline then progressively repositions across the bench, excavating overburden and exposing coal as it goes

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A. Starting point

Cast to Final

A

A'

B. Cast blast and bench leveled

A'

A

C. Excavation commenced and bench extended

A' Rehandle

A

D. Excavation completed and coal edge cleared

Figure 10.8-9 Range diagram sequence

(the block). After it has uncovered the full strip width of coal, it repositions diagonally down the strip into position to commence digging the next key. The method has higher rehandle but is productive because of the short swings, the previous strip floor can be cleaned up prior to overburden being placed in it, the highwall is dug in-line giving a safe wall, the coal edge is also excavated in-line, and it is a simple method to execute. A variant of this, called the extended key with inpit bench method (Figure 10.8-10C), involves the key being excavated for 10 or 20 blocks and an in-pit bench built on the low-wall side. At the end of the extended key, the dragline then repositions to the in-pit bench and “pulls” the remaining blocks for the length of the extended key, exposing coal. An advantage of this method is lower rehandle than the key bridge, as the in-pit bench does not have to be filled in completely from the highwall to the low-wall side (no “bridge” required). In the chop cut with in-pit bench method (Figure 10.8‑10D), use is made of throw or cast blasting to move as much of the

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2

1 2

A. Side cast

3

1

D. Chop cut with in-pit bench

2

3

1

B. Key bridge

2

1

E. Double in-pit bench (elevated bench)

Pullback Step 1

2

1

C. Extended key with in-pit bench

1

2

F. Pullback

2 3 = operating positions for the dragline throughout the dig sequence Gray areas = where material is dug from Black areas = where material is dumped to

Figure 10.8-10 Main dragline methods

prime material across to the low wall as possible. The dragline excavates the key area along the strip by digging perpendicular to the new highwall and creates an in-pit bench toward the low-wall side of the strip. The dragline then retreats along the in-pit bench and excavates the block that it was previously sitting on and places it in final position. The method has low rehandle and high prime productivity but is a more complicated two-pass operation that gives poor control of highwall. The double in-pit bench method (Figure 10.8-10E) is used when the depth of material is such that the dragline operating envelope is not sufficient to place all of the overburden into a final position from a single working level. For this reason it may also be known as an elevated bench method. An extended key or chop cut with in-pit bench method is used, but instead of the bench being fully excavated to final position, a partial excavation, or trim, is undertaken and the overburden placed to create another bench at a higher working level. When a suitable length of upper bench is created, the dragline moves onto the upper bench and excavates the remaining waste to expose the full strip width of coal. The higher second bench increases the volumetric capacity of the dragline and so allows

the exposure of deeper coal. Walking delays increase, overall productivity is lower, the method is more complex, and coal exposure is more intermittent. Although not a specific method, the pullback technique (Figure 10.8-10F) involves walking the dragline up into the spoil to pull existing overburden back prior to placing overburden from the current strip. It is time-consuming to prepare access for the dragline, and the pullback material does not expose coal so it is all rehandle, but productivity can be reasonable depending on the volume to be excavated. It is rarely used as part of an ongoing operation because of its high cost, but it can be used in areas where spoil room is particularly tight, such as around ramps or deep boxcuts or to recover from design errors or geotechnical failures. Operationally this method can be delivered by two separate draglines: one on spoil in the pullback mode and one on the bench to be mined. As the seams progress to greater depths, typically at around 85 m, a truck and loader prestrip operation must be introduced to assist the dragline to continue to reach the top of the coal by creating a working level for the dragline that is lower than the natural topography. Prestrip may have

Strip Mining

been introduced sooner than this, however, depending on the deposit and dragline geometry and general economics. Prestrip may be directed to relieve the dragline in tight spoiling areas such as when crossing ramps or endwalls or where the natural topography is variable, or prestrip may be applied more generally across the whole strip. Case Study: Norwich Park Mine

The Norwich Park hard-coking coal opencut mine is located on the western fringe of the Bowen Basin coalfield in Central Queensland, Australia. The economic seams are contained in the Mid- to Late Permian German Creek Formation, which are overlain by up to 55 m of poorly consolidated cemented sediments consisting predominantly of sand and clay with irregular gravel beds and weathered basalt flows. The depth of weathering varies from 15 to 25 m north of Leichhardt pit to 25–50 m in the Leichhardt pit and south. Four coal seam groups are present over the mining area with the dip of the sequence to the east and varying from 2° to 10° in some areas. Equipment selection—Physical and volumetric scale of extraction. Initial mining operations commenced on the subcrop of the Dysart seam(s) in 1979 to uncover coal in strips oriented along the strike of seams. Mining has progressed along the strike and downdip over the intervening period. Surface features are generally flat with a few ephemeral creek systems running across the deposit. The mining area has been divided into a number of pits with these features incorporated into the layout design. The mine processing plant and other facilities are located on the western side of the deposit in a generally southern location. Currently, the mine employs a fleet of major mining equipment consisting of six electric walking draglines supported by excavators and rear-dump trucks for waste removal and front-end loaders and bottom-dump trucks for coal mining and haulage. The coal haulers haul ROM coal to the crushing and processing plant along a haul road network. Total annual product metric tons are typically 6 Mt from a plant feed of 8.5 Mt running at an average yield of approximately 71%. Total annual prime overburden moved is typically 100 Mbm3 giving a prime-to-product strip ratio of approximately 17:1. Allocation of this prime is mainly to dragline (60 Mbm3/yr) followed by prestripping (27 Mbm3/yr) and the remainder to production dozing and minor waste operations. Operating methodology. Figure 10.8-11 provides a schematic of the mining process representing the typical dragline and truck and shovel activities. Truck and shovel stripping is used to complement and expedite dragline productivity with waste hauled around the dragline operation and dumped on top of dragline spoil. Typically, the truck dump is two spoil peaks behind the current dragline strip. Single-seam dragline methods are used in the Price and Leichhardt pits, where successive seams are uncovered using trucks and excavators, as well as areas of Campbell and Roper where the upper seams are coked or weathered. The two methods used in the single-seam areas can be broadly categorized as the Jensen and Curtis off-line key bridge method (conventional or extended key), and the extended key and elevated bench method. The conventional Jensen and Curtis method progressively uncovers the full seam width, whereas the extended key does not. The extended key method is used where the height of the elevated bench above the in-pit bench

1003

is too great to use the conventional Jensen and Curtis method. This might be the result of an area that has a very low in-pit bench because of the blasting and dozing techniques, or an area where the elevated bench has to be built high to enable the dragline to exit the pit. The latter is often used from the endwall back and is known as a reverse key. Little difference exists in the planning requirements in regard to range diagram analysis for spoil fit for these two methods. The main difference relates to the timing of the coal exposure. The conventional Jensen and Curtis method has continuous coal exposure whereas the extended key method has sporadic coal exposure due to the need to return to the front of the strip to pull blocks to expose the full seam width. Jensen and Curtis off-line key bridge method— Leichhardt pit example. After blasting and any dragline prestrip, a dragline access road is formed to the ramp mouth. If time permits, dozers will be used to bulk push waste across the pit and lower the in-pit bench level. The dragline will then be walked to the ramp mouth and the dozer key will be set up to break away from the ramp mouth. This initial waste is spoiled behind the dragline to form the in-pit, or chop, bench. The dragline will then be positioned on the chop line to start working away from the ramp. The dragline digs key material and dumps it into the in-pit bench. The dozer will be cleaning coal and pushing up spoil on the low-wall side at this stage. The dragline will widen, or trim the key, and throw the material to final spoil. The dozer will continue preparing the in-pit bench. The dragline works off the in-pit bench and throws the remaining low-wall side material to final spoil. At this stage the dozer will be working against the highwall, cutting down the key material and cleaning the highwall. For the rest of the strip, dozer push will be used to excavate the highwall portion of the key and expose down to the top of the coal. The progress of the method in sectional view is identical to the example range diagram sequence provided in Figure 10.8-12. Operationally, the method makes extensive use of dozer operations to cost-effectively supplement the dragline operation. Extended key and elevated bench method—Roper pit example. Using this technique, the floor of the previous strip is cleaned before blasting. Any mud and water is dammed up on the low-wall side. The area is not presplit before blasting. The overburden material is cast blast onto the pit floor with the blastholes drilled at 75°. After blasting, the dragline takes a highwall trim from the surface (Figure 10.8-12A) down to 20 m below the surface and forms a safety bench. The dragline or dozers then prepare an access into the pit and will prepare the chop bench. The dragline will walk into the pit and commence digging off-line extended keys from the ramp mouth (Figure 10.8‑12B). The spoil will be used to build an elevated bench. When coal needs to be exposed to suit coal-mining requirements, the dragline will form a ramp up onto the elevated bench and will walk back to the front of the strip to pull blocks, exposing the full width of the Dysart seam (Figure 10.8-12C). Additional low-wall trim will be taken to ensure a stable lowwall angle and a defined low-wall edge offset. Truck and shovel prestrip. Prestrip is usually conducted by contract. Planned prestrip is the difference between the waste removal requirement to uncover the annual coal production and the installed dragline capacity. Shovel/truck priorities include geotechnical recommendations, spoil relief, and poststrip activities. Free-dig prestrip areas comprise horizons that cannot be drilled because blastholes will not stay open (i.e., they are too sandy).

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Meters +100

Dragline Dragline Spoil

Truck & Excavator

0

Harrow Creek Lower and Upper Seams

P Seams

Base of Weathering

Rider Seam

–100

Dysart Seam

–200 0

100

200

Meters

Figure 10.8-11 Cross section of Norwich Park mine Variants at Other Mining Operations

A number of possible variants of single-seam dragline operations exist. Track dozers may be deployed to assist the dragline operation and speed up coal exposure rates as at the nearby Gregory mine. This mine has also used excavator and truck fleets to remove the dragline key material to accelerate coal exposure. The advantage is that this material is placed outside of the dragline operating area and does not contribute to rehandle, unlike production dozing. Shallow mines in the United States have employed a direct-cast or side-cast method. The operating strip width is narrower and the wasteto-coal ratio is such that the dragline is able to directly place all material into the final position without building working pads and thus incurring rehandle. For steeply dipping seams or highly faulted deposits, the dragline may expose a “false floor” and leave the remaining wedge of material overlying coal to be removed by smaller truck and excavator methods. Multiple-Seam Dragline Operation with Prestrip For multiple coal seams that occur deeper than 30 m, and where economies of scale allow, a multiple-seam pit configuration is selected. Within this broad configuration a combination of a number of the single-seam dragline methods is applied to uncover each successive seam. Strip-and-block geometry is similar to single-seam dragline methods, but the geometry dictates significant time and effort to move the dragline up and down between what can be quite different working horizons. A brief overview of the two broad categories of multiple-seam pit configurations—namely, stacked and offset—follows. Stacked Configuration

In a typical stacked configuration, the dragline removes the overburden in a sequential operation working from the top of the sequence to the bottom. Multiple dragline passes are required with one coal seam being exposed on each pass. Generally, the first pass is a simple direct-cast operation into the void from the previous strip. Subsequent passes will generally involve key or chop operations and extended benching to expose the coal and place the spoil into its final position. The stacked configuration can result in low rehandle (25% or less) and high productivity, especially for narrow strips. One disadvantage is that the lower-pass burden blasts are buffered by upper-pass overburden, and so less cast blast is achieved and fragmentation can be poorer, leading to reduced digging

A. Highwall trim

B. Extended key and elevated bench

C. Final blocks

Figure 10.8-12 Extended key and elevated bench sequence

productivity. A second disadvantage is that the material from the upper pass ends up on the floor of the whole spoil pile, may not be suitable to use as a spoil base, and can frequently remain in the base of the pit for extended time between passes where it undergoes further degradation. A third disadvantage is that if lower passes are relatively thin compared to the upper pass, high rehandles can result. Offset Configuration

In a typical offset configuration, the upper seam is offset at least one complete pit width from the lower seam. This allows

Strip Mining

the lower pass to be blasted first, followed by the upper pass, which is generally cast as much as possible to lower the dragline working level. The overall sequence is then very similar to an extended bench method where the extended bench covers the entire lower seam. It is important that the bench level is as low as possible as the bench is built entirely from rehandle. Advantages include the exposure of two coal seams simultaneously, which allows them to be mined at the same time enabling blending. Blast performance is improved for the lower pass as it is not buffered, and scheduling is simplified as the dragline does not require an alternative working area while a seam is mined out and the next pass of overburden is drilled and blasted. Disadvantages are that overall rehandle is generally higher and can increase dramatically with increased total waste allocated to the dragline, and in deeply prestripped mines the large open void requires a significant investment in working capital to establish and maintain as well as placing further stress on the dump volume balance. The selection of one configuration over the other is dependent on an analysis of many factors: • Seam dip • Relative geometry of the upper- and lower-pass overburden and seam thicknesses • Geotechnical characteristics of the overburden • Scheduling and blending considerations Case Study: Goonyella Riverside Mine

The Goonyella Riverside high-quality hard-coking coal opencut mine is located on the western fringe of the Bowen Basin coalfield in Central Queensland. The mine area has a strike length >30 km arranged in two groups of parallel pits representing the originally separate Riverside and Goonyella mines. In the west, pits target the single basal Goonyella lower seam (GLS); in the northeast, pits target all three major seams: the Goonyella middle seam (GMS), Goonyella upper seam (GUS), and the GLS; and in the southeast, pits target the GMS. An underground longwall operation is currently installed in one of these pits and is extracting GMS coal from panels oriented perpendicular to the strike. Pit configuration—Geological orientation. The Goonyella deposit contains economic seams in the Late Permian Moranbah coal measures that are approximately 300-m thick. The Moranbah coal measures are Mid- to Late Permian age and are characterized by several laterally persistent, relatively thick coal seams interspersed with several thin minor seams. The Permian coal measures dip east at 3°–6°. The mine area is covered by 0.5–30 m of poorly consolidated Cainozoic sediments consisting of lenses of river channel gravels and sands separated by sandy silts, sandy clays, and clays. The Tertiary silts and clays are densely compacted, hard, and generally dry. The major seams commonly exhibit decreasing ash content and increasing vitrinite content toward their base and are recognized for their superior coking properties. Equipment selection—Physical and volumetric scale of extraction. Mining operations commenced on the subcrop of the GMS in 1971 to uncover coal in strips oriented along the strike of the seam. In 1983, mining of the GLS commenced following the commissioning of the Riverside mine updip of the GMS operation. Also in 1983, the Goonyella mine commenced mining the GLS in addition to the GMS in a doubleseam operation.

1005

Surface features are generally flat with a few ephemeral creek systems and the Isaac River traversing the southern portion of the mining lease. The mining area has been divided into a number of pits with these features incorporated into the layout design. Currently, the mine employs a major mining equipment fleet consisting of seven electric walking draglines supported by a number of electric rope shovels, hydraulic excavators, and rear-dump trucks to effect waste stripping. Coal mining is undertaken by front-end loaders and rear-dump trucks. The trucks haul ROM coal to one of two crushing and processing plants along a haul road network. Total annual product metric tons are typically 16 Mt from a plant feed of 20 Mt running at an average yield of approximately 80%. Total prime overburden moved is typically 125 Mbm3/yr giving a prime-toproduct strip ratio of approximately 10:1 for the opencut coal component. Allocation of this prime is mainly to truck and shovel (80 Mbm3/yr) followed by dragline (45 Mbm3/yr). Operating methodology. Figure 10.8-13 provides a schematic of the mining process representing the typical dragline and truck and shovel activities. Truck and shovel waste is hauled around the dragline operation and dumped on top of dragline spoil. Typically, the truck dump is two spoil peaks behind the current dragline strip. The four northern double-seam pits target the GMS and GLS and account for 70% of opencut production. The GUS is being exposed and mined in the prestrip areas of these pits. The double-seam pits in the north have the lowest strip ratio, and production is concentrated in this area. All double-seam pits have low-wall access ramps. The remaining single-seam pits target the GLS in the west and the GMS in the east and account for the remaining 30% of opencut production. As the single-seam GMS pits advance downdip, they are picking up the GUS in their prestrip operations. The choice of dragline technique depends on a number of pit-specific conditions. Several dragline techniques are used that can be broadly categorized as • • • •

Single-seam key bridge methods, Single-seam key bench methods, Double-seam offset methods, and Multiple-seam stacked methods.

Although the techniques are long-established for each pit, the basic selection criteria for single-seam methods is whether the low wall is geotechnically stable, with an extended key/ bench method used where it is and a key bridge method used where it is not. In the dual-seam areas, the offset method is favored unless the interburden is considerably thinner than the overburden. Double-seam offset method—Redhill pit example. The first step is to blast the interburden, which is cast blast followed by blasting of the overburden with material spilling over the interburden bench. The dozers then form a working pad on the shot overburden. Sitting on an off-line position, the dragline will then uncover the full width of the GMS. This essentially consists of a combination of off-line key and trim (widening) of that key to expose full-seam GMS. The spoil is cast beyond the GMS coal edge. Ideally, the entire length of GMS in the strip will be uncovered before the dragline has to move onto the in-pit bench. A combination of selective placement with the dragline and use of dozers creates an in-pit bench. The dragline forms and walks across a bridge onto the in-pit bench. This bridge is subsequently removed.

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Meters

Prestrip Truck Dump

+100

Dragline 0

Goonyella Middle Seam

–100

Goonyella Lower Sea m 0

–200

Truck & Shovel

Dragline Spoil

100

200

Meters

Figure 10.8-13 Cross section of Goonyella Riverside mine

The operation is now the same as a conventional off-line key/ elevated bench method. From an off-line position, an extended GLS key and trim is taken, and the spoil is used to build an elevated bench. The dragline walks up onto the elevated bench and pulls blocks to uncover the full width of GLS coal. The GMS coal is accessed from the highwall whereas the GLS coal is accessed by a low-wall ramp at the southern end of the strip. Multiple-seam stacked method—Thiess pit example. The upper split of the GLS (GLUS) overburden is blasted and a working pad is formed on top of the blasted ground. Full-width GLUS is uncovered by taking an on-line key, then trimming (widening) the key until the GLUS is exposed. This is essentially a side-casting operation. After the GLUS is exposed, it is accessed via a temporary coal ramp that is later removed by the dragline. Then the interburden is drilled and fired. The operation is now the same as a conventional off-line key/elevated bench method. From an off-line position, an extended GLS key and trim is taken, and the spoil is used to build an elevated bench. The GLS blocks are pulled from the elevated bench to expose the full strip width of the lower seam coal. The GLS is accessed by a conventional center ramp. Truck and Shovel Operation As previously discussed, truck and shovel strip mining is selected where flexibility beyond that offered by an efficient dragline method is required. Typically, these applications are in more complex multiseam and/or steeply dipping deposits. Truck and shovel is preferred in steeply dipping deposits where waste spoil room in the previous strip is insufficient within the efficient operating envelope of a dragline. Truck and shovel is also used for shorter life operations where there is insufficient mine life to provide a payback for the higher capital investment of draglines. The two main types of truck and shovel mining operations are conventional along-strike operations, with the pit progressing downdip with each successive strip, and downdip mining or terrace mining, where the strips are excavated downdip usually to the final highwall and then progressed along the strike (effectively at right angles to a conventional along-strike method). The advantage of terrace mining is that the waste material can be short hauled either across the blasted waste or around the ends of the strips to the adjacent waste dump. Typical dimensions for terrace mining in Australian coal mines are 200 m downdip using 70-m strip widths. Terrace mining is very attractive for steeply dipping

coal seams where the steep floor conditions significantly reduce the spoil room available for conventional along-strike methods. Examples of terrace mining include Australian coal mines such as Macarthur Coal’s Moorvale mine, Peabody’s Burton mine, Jellinbah Resources’ Jellinbah East mine, and Cockatoo Coal’s Baralaba mine. Case Study: Mt. Arthur North Mine

The Mt. Arthur North mine is an example of a conventional along-strike truck and shovel operation. It is located in the Upper Hunter Valley of New South Wales, Australia. The mine site has a topography that is moderately undulating, steepening near the base of Mt. Arthur. Domestic product is transported via an overland conveyor connecting the mine to the Bayswater Power Station, and export product is shipped from Newcastle. Pit configuration—Geological orientation. Coal seams within Mt. Arthur Coal’s leases are contained within the Wittingham and Greta coal measures, which are separated by a thick noncoal-bearing unit known as the Maitland Group. Coal seams split and coalesce throughout the area, with 20 discrete seams splitting to some 80 unique coal plies. In the northern part of the deposit, the seams dip from subcrop toward a major monocline at dips of 5° to 10°. Across the monocline, seam dip increases to 10o–25o before flattening again to dips of 2° to 4° in the broad Calool syncline. Indications are that the seams roll over a broad and less-welldefined Denman anticline west of the Calool syncline. To the west of the Denman anticline, seams are down-thrown by the Mt. Ogilvie fault zone. The in-situ waste strip ratio (cubic meters per metric ton) generally increases to the southwest with current opencut planning targeting areas of up to a 5:1 strip ratio, which extend downdip to the edge of the eastern monocline. Beyond the monocline, the dip steepens locally and the basal seam plunges to depths >400 m before the seams flatten out. Equipment selection—Physical and volumetric scale of extraction. The mine is currently capable of operating at a nominal 14.5 Mt/yr, based on 73.5 Mm3/yr waste prestripping capacity (including contractors) at a 5.1 m3/t strip ratio. The choice of mining equipment was dictated by equipment operating at the adjacent Bayswater No. 3 mine—electric rope shovels and hydraulic backhoes, which were progressively transferred to Mt. Arthur North during 2001–2002. Although the rope shovels provided a proven low-cost prestrip capability, their inability to operate on steep dips and their

Strip Mining

limitations in shallower interburden resulted in the procurement of additional hydraulic backhoes to mine the “wedge” of overburden that remains beneath the rope shovel floor and the next coal seam as well as interburdens in the 2-to-10-mthickness range. The rope shovels are generally limited to the “base-of-weathering” pass and the thicker prestrip passes. The current distribution of prestrip by operation type is 35% by rope shovel, 45% by large hydraulic excavator, 10% by smaller hydraulic excavator (coal/parting fleet), and 10% by mining contractor. An 80-m strip width was chosen to maximize the use of two electric rope shovels while providing sufficient width to operate. Overburden is blasted prior to excavation. Coal is mined using smaller hydraulic excavators and a front-end loader. Coal is “free-dug” in situ. Very thin coal seams may be ripped by a bulldozer and pushed into piles for more efficient loading. Operating methodology. The mining method is open-pit strip mining, using truck and shovel methodology, spread over multiple prestrip horizons, as shown in Figures 10.8-14 and 10.8-15. The advantages of this system are low technical risk, high resource recovery, and flexibility to change sequencing and production rates as required. The pit design is based on a strip mining layout where the strips are oriented parallel to the strike. Having fully developed along a 6-km strike length, the mine progresses downdip in 80-m-wide strips. The coal seams dip at between 5° and 8° over the majority of the lease; however, the dip increases in the far southeastern corner to 15°–20° where a bench mining technique and through-seam blasting is used. In the main pit there are large areas where the seams dip <7°, allowing hydraulic excavators and rear-dump trucks to work on grade. Thicker interburdens are dug by rope shovels and hydraulic excavators. The rope shovel is used to prestrip the majority of the interburden, leaving a remnant wedge for a backhoe to remove. Above 7° dip, it is necessary to reduce the apparent dip and work off temporary benches. In these steep dip areas, track dozers are required to prepare work area pads and/or push down and stockpile material ready for loading. Material to be excavated is first prepared by blasting or ripping, or it is free-dug. Waste rock is hauled from the highwall to the low-wall side of the mining void to backfill mined-out areas. To cross the mining void, truck haulage is either along the mining strips and via the pit endwalls or via highwall ramps and low-level cross-pit bridges that connect the highwall to the low wall. Although generally a shorter distance, the cross-pit bridge system may necessitate significant de-elevation and re-elevation of material. Most mining blocks will have a choice of the two types of routes, and the one that offers the quicker cycle time is the one that will be selected. Shovel and RearDump Trucks for Overburden Rip and Push Thin Coal Seams

Figure 10.8-15 Mt. Arthur mining method

1007

A high-level cross-pit bridge is installed between the main pit and the Ayredale pit and functions as an endwall. ROM coal is mined in situ using hydraulic excavators in both backhoe and shovel configuration, according to the working section thickness. Coal is hauled to the ROM coal receiver hoppers for delivery to the coal handling and processing plant. ROM coal can also be stockpiled nearby to the receiver hoppers and rehandled later as required. Coarse reject from the coal washery is conveyed to an overhead truck loading bin, situated adjacent to the coal receiver hoppers and backhauled by trucks into the spoil dumps on the return haul to the coaling face. Spoil dumping has migrated from ex-pit dumps during the start of the operation toward in-pit placement. This involves dumping on sloping floors up to 9° and generating a layback of the dump profile for stability. The layback is enhanced by a low-wall ramp and haulage system incorporated into the dump face. Nominal batter heights of 20 m and bench widths of 40 m are used for initial dump design and suitable for standard truck-tipping operations at the crest. This profile of 2:1 is able to be locally optimized to larger batter heights and smaller bench widths where floor conditions are deemed suitable and material composition is homogeneous and competent. Floor preparation standards are enforced on the Ramrod Creek coal seam floor. This normally involves ripping and removal of silt, water, and remnant coal. Floor blasting has not been required but will be evaluated for future steeper dip areas. Placement of material in the dump is managed, and standards exist for the type of material to be placed on floors or high tip faces. Coal reject and base of weathering material

Figure 10.8-14 Mt. Arthur coal pit operations

Reshaping Wedge Removal

Coal Mining

Dumping

Topsoiling and Planting

1008

Table 10.8-1 Strip mining strategies compared by selection driver Strip Mining Strategies Subcriteria

Driver

Cast, Doze, Excavate

Dragline

Truck and Shovel

Geology Complex geology of ore–waste interface impacts, ore recovery, and equipment productivity. Selective mining practices are required.

Small scale of equipment units enables effective operation.

Ancillary equipment required for ore–waste interface. Large draglines can result in significant ore losses.

Flexibility of equipment enables resource recovery but with slower loading rates and some impact on costs.

Material strength

Instability due to low bearing capacity of soft waste material

Dozer production is only slightly impacted due to low equipment ground-bearing pressure. For extreme cases, install wider dozer tracks.

Dragline can sink into soft material. For localized weak materials, dragline can be used to dig off soft material and replace with competent. For more extensive occurrences, a buttress needs to be maintained to provide stable ground for the dragline. Equipment modifications include larger tub and wider shoes.

Shovel is also prone to sinking in unstable material. Equipment modifications include significantly wider tracks. For shallow soft waste, the shovel can be operated on the stable floor below the soft material, and if necessary, dozers can lower the material to a workable bench height.

Slope stability

Unstable material results in flatter highwall and low-wall batters and less-stable working faces.

Negligible impact due to the normal dozer operation having relatively flat highwall and low walls

Flatter working slopes require longer boom and smaller Fairly insensitive to working slopes with minor bucket dragline configurations to provide the required increase in haulage distances resulting in very reach resulting in higher unit costs. minor increases in unit costs.

Depth

Waste-stripping cost as a proportion of total mine costs

Optimal cost for shallow depths 40 m or less; deeper than this, dozers become less productive with long and uphill pushes.

Optimal cost for 40–85 m, may need some prestrip assistance; deeper than this requires a significant waste-stripping operation above the dragline.

Optimal cost for waste above 85 m.

Dip

Geotechnical stability and deposit geometry starts to favor certain equipment characteristics.

Suitable for flat to moderate dips; at steep dips, need to cast blast to a false floor and pick through rest of waste to expose coal with secondary truck and excavator operation.

Best for flat to moderate dips; at steep dips, low-wall stability can be an issue and capacity of dragline to reach top of coal and place the overburden in a final position is severely limited. Can expose a false floor and follow up with excavator and trucks.

Ultimate flexibility for all dips. For very steep dips, excavators will pick through waste and coal and may need to load trucks on the same level the excavator is sitting on; less productive.

Seam thickness

Impact of economics for thick seams and driver to selective mining

Ideal for thin seams where dozers and excavators can expose the coal with minimal loss; tool is selective and operator is close to the coal–waste interface.

Best for seams 1 m and greater; thin seams can be severely disrupted by large blasts, and bucket is a large digging tool trying to follow a waste; coal interface many meters away from the operator; excessive losses an outcome for thin seams. Thick seams will be profitable if mine is able to fund capitalintensive draglines.

A thick seam can subcrop economically at a depth beyond the reach of a dragline; leads to deep truck and shovel boxcut; potentially stick with truck and shovel due to impracticalities of getting dragline in and out of a very deep pit.

Number of seams

Waste geometry between seams drives selection of most appropriate method.

Overburden pushed into the low wall needs to end up in a final spoil position away from the coal to be mined; many seams will result in dozed spoil rilling back against the coal edge; will lead to loss or dilution.

Best operating on a few relatively thick waste sections; intensive set-up effort to get a dragline to ramp down into a pit and out again and will prefer to have a reasonable task when it is set up; prefer interburdens of minimum 10-m thick. Multiple seams at the bottom of the pit with thin interburdens mean the dragline will not be able to effectively clear the coal edge, leaving spoil up against it; creates a truck and excavator task to excavate the low wall back.

Best for many thin seams especially where waste between seams is thin or seams are split; is able to take the waste completely away from the coal, leaving it to be recovered cleanly.

Deposit

(continues)


Faulting

Table 10.8-1 Strip mining strategies compared by selection driver (continued) Strip Mining Strategies Driver


Dragline

Truck and Shovel

Stratigraphy development

Impact of upper seams subcropping at depth

This method is generally limited by depth; however, if upper seams subcrop, they interfere with the effectiveness of the cast operation, with thin overburdens difficult to cast and upper overburdens potentially buttressing lower interburdens and reducing overall cast effectiveness.

As more upper seams enter the sequence, the whole mine, driven by the downdip advance of the dragline, will actually generate more coal; one option is to expand output, but if this is not economical, the choice is either to lift draglines up in the sequence and abandon lower seams, raise dragline level to reduce advance (give more waste to the dragline), or, if it is a multiple dragline mine, to stand some draglines down as no longer required as the mine transitions to source more of its coal from the overlying truck and shovel pit.

Working multiple subcrops at once can lead to spoil from the downdip pit being placed on top of future potential resource as the updip pit catches up; truck and shovel systems provide the flexibility to place overburden more selectively to allow updip pits to mine through more effectively.

Size

Resource size will determine the ultimate scale of the mine to be constructed.

Optimal for small deposits that cannot justify large capital expenditures with long payback periods

Optimal for large long-life deposits seeking to generate Highly scalable equipment that can be selected large volumes for a long period of time where capital to match the deposit, both initially and also as it payback occurs. Comes in large lumps of capacity, so develops over time. potentially need to supplement with truck and shovel stripping capacity. Ends up as either prestrip or a satellite truck and shovel pit only.

Topography

May influence choice of equipment and therefore mining method

Will affect balance between cast, doze, and excavate.

Highly variable topography is problematic for dragline; will generally lead to supplementary truck and shovel stripping.

This is the most flexible mining system, but truck grade limitations and variable topography can result in waste being brought forward in the schedule to access upper benches. Dumps can also lose capacity if topography is falling away on the dump side.

Geometry

Physical constraints to mining operations will influence selection of equipment and therefore mining method.

Highly flexible and well suited to variable deposit geometries

The preference is long, straight strips; short strips lead to lots of walk time and reduced productivity. Bends in strips can lead to inside corners that create spoil tight spots; this can be alleviated with truck and shovel assistance. It is very difficult to extend the strip after the mine has been developed to depth, essentially trying to dig a deep boxcut; this generally needs to be done with a truck and shovel. Vertical development to pick up deeper seams is a similar problem; a truck and shovel is needed to develop.

A coal seam that subcrops at the lease boundary means that the boxcut may need to be constructed by truck and shovel and the waste hauled elsewhere. Multiple valleys through low-wall dump mass means that for pits deeper than 150 m, total dump balance will come under stress; this may favor a move to hauling to highwall side and other out-of-pit locations; or for really long hauls, a conveyor waste system may be more economical.

Location

Lack of supporting infrastructure may influence choice of equipment and therefore mining method.

Smallest equipment and easiest to support from an infrastructure point of view

A lack of availability of electrical power, in total or reliable supply, may preclude use of draglines. It may be difficult to bring large equipment to location. A lack of specialist operating labor and technical support labor may also detract from draglines.

A lack of electrical power may preclude use of electric rope shovels and lean toward diesel over hydraulic excavators. A lack of specialist technical support labor may also drive toward smaller equipment.

Strip Mining

Subcriteria

(continues)

1009

1010

Table 10.8-1 Strip mining strategies compared by selection driver (continued) Strip Mining Strategies Subcriteria

Driver


Dragline

Truck and Shovel

Environment Noise limitations on operation due to proximity of third parties

Tracks and engine revolution are main sources, but activity is mostly in-pit, hence less off-site impact.

Significant noise from bucket and rigging. Dragline regularly dumping above natural surface and can carry off-site. At worst, may exclude draglines from sensitive areas close to third parties.

Significant noise from engine revolutions on acceleration and tipping, horn signals, and reversing alarms. Trucks are often dumping at height and locations remote from the active pit near the mine boundary. This can be reduced by technology (“whisper quiet” trucks). For sensitive areas, a limit on traffic intensity and hours of operation can be required.

Dust

Proximity to third parties

Minimal as mostly slower material speed and restricted to in-pit, therefore less impact

Significant due to dumping at height driven by dragline geometry

Significant with dust generated at the load point, haul roads, and final dump. Dust is usually controlled by adequate work area and haul road watering. Controlling dust generation from large elevated dumps is a challenge.

Economics Capital

Capital cost per bank cubic meter

Lowest with shortest lead times due to factory line production of equipment

Highest and longest lead times due to single unit fabrication and on-site erection

Mid-range with variable lead times (1–3 years) depending on supply–demand balance

Labor

Total cost of labor and accommodation per bank cubic meter

Mid-range with typically 10 operators per operating shift for 15–20 Mbm3/yr capacity

Lowest with typically 2–3 operators per operating shift for 15–20 Mbm3/yr capacity

Highest with typically 15 operators per operating shift for 15–20 Mbm3/yr capacity

Other costs

All other operating costs per bank cubic meter

Mid-range

Lowest with largest cost driver being electricity

Highest with largest cost drivers being fuel, tires, and maintenance

Expansion capability

Ability to increase production incrementally

Additional units add small capacity steps.

Minor step increases possible via motor upgrades; beyond this, additional units provide large-capacity steps.

Additional capacity by adding trucks up to loader; beyond this, additional units provide large-capacity steps.


Noise

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Table 10.8-2 Forces for change and their impact on strip mining Fundamental Force Health and Safety Focus • Regulators demand safer mines • Shareholders demand safer mines • Mine operators and mine workers demand safer mines

Environmental Concern • Global warming • Climate change • Land degradation • Biodiversity

Social Forces • Urbanization and aging work force • Environmental values

Technology • Automation • Increased equipment scale

Drivers for Change

Potential Outcomes

• Continuous reduction in injury rates demanded by all • More research reveals adverse long-term health outcomes

• “Smart” mining equipment and maintenance tools detect

• Public perception of greenhouse gas–intensive companies

• Time-based Google Earth with enhanced image-processing

• Increasing service gap between metropolitan and regional

• More-prevalent fly-in/fly-out as the standard model • Number of personnel on-site an absolute minimum • Real-time terrain acquisition means most geology, survey,

from shift work and other mining-related activity • More risk of injury for aging work force

and their products; remoteness of most mines does not deter environmental action groups from focusing public attention • Carbon taxes and trading • Water scarcity with users competing for allocations (domestic, industrial, and agricultural); mine discharge water quality becomes higher profile as background or natural flows reduce • Food security becomes topical with opencut mining targeted as an activity that degrades prime agricultural land

or remote areas; harder to attract highly skilled personnel to remote areas • Societal pressure for increased external control of licenses and leases (the social license)

unsafe practices and require third-party risk assessment and lockout override before proceeding • “Smart” personal monitoring system detects unsafe behavior and raises central alarms and automatic notification to wearer and supervisor • Remotely operated, semi- or fully autonomous equipment with robots undertaking routine maintenance

tools means anyone on the Internet can monitor net land disturbance and rehabilitation deficits, leading to more focus on progressive rehabilitation • Electricity generated by nuclear power potentially tips economics toward electrically driven equipment, leading to emergence of widespread use of electric excavators and electric trucks using overhead cables; alternatively, more use of conveyors to move waste • Water collected from across the whole mine site at closure and directed into final voids that have been left open as large catchment dams, leading to treated water sold to local users • Underground mining becomes only form of mining allowed in areas of, or near, high-quality agricultural land • Carbon taxes tip economics heavily in favor of underground mining due to much-reduced energy expenditure

engineering, as well as equipment performance and condition-monitoring tasks occur off-site in centralized technical centers • Region must be persuaded of sustainable development before it awards the social license

• Need to eliminate safety incidents • High cost of labor and associated costs including accom-

• Less human exposure to accidents • Fully automated operations: truck–shovel fleets, draglines,

• Increasing strip ratios for deeper strip mines • Shortage of suitable experienced personnel

• Reduction in on-site labor requirement to maintenance and

modation

and production dozing control functions

• Significantly higher productivity • Lower maintenance costs as more predictable operating

practices

• Outcomes would have the largest impact on truck and

shovel costs

Market Forces • Continued globalization • China/India industrialization

• Vertical integration for commodity security; commodity

• Marginal undeveloped deposits become economical; thinconsumers buy and develop mines ner seams, so more likely to be truck-based mines • Supply-side consolidation; companies continue to aggre• Existing deposits economical at greater depths gate to pursue improved economics • Huge variation in prices in short-term timeframes means • Long-run prices reflect relative scarcity of energy and metalthere are no final highwalls; major mining companies exit lurgical coal in the face of sustained increasing demand but mines and they are purchased by equity hedge funds tradwith short-run volume and price volatility ing on cash generated between care and maintenance and incremental marginal extraction cycle • China becomes a major manufacturer of draglines and earthmoving equipment; much reduced-capital pricing fundamentally shifts relative economics of equipment

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must be incorporated into the dump tip faces and blended with stronger material to minimize their negative stability impact.

STRIP MINING STRATEGIES COMPARED BY SELECTION DRIVER

Having outlined the four major strip mining strategies, it is useful to explore the relative suitability of each strategy in a given mining application. Four significant decision criteria have been identified: geology, the deposit, environmental considerations, and economics. For each of these, a number of subcriteria and their respective drivers and influences on each strip mining strategy are summarized in Table 10.8-1.

FUTURE TRENDS

This chapter has outlined a range of pit configurations, equipment types, and operating methodologies that represent the breadth of current strip mining practices. Although the specifics of the future are uncertain, one thing is definite—change comes. In closing out this topic, Table 10.8-2 identifies some of the major forces at work today, comments on the drivers for change they introduce, and speculates on their potential outcomes. The past 100 years in strip mining has seen the emergence and continuous development of powered earthmoving equipment. Today’s trucks move >300 t/cycle and are filled by large hydraulic excavators or rope shovels in as few as three passes. Electric walking draglines have a dig-to-dump range

>100 vertical meters, a similar operating radius, and move more than 200 t per 1-minute cycle. Advances have primarily come through scale (enabled by improved materials and engineering design), application of low-cost and mechanically reliable electric motors, and technology generally for improved mine planning, equipment operation, and condition monitoring. The result has been existing strip mines that remain economical at depth and previously uneconomical deposits now within reach. Recent rapid development in the areas of geospatial awareness, image processing, real-time communications, and general computing power suggest that the industry is on the brink of significant breakthroughs in remote and autonomous equipment and even whole-of-mine operation. A step change in the relative economics of conveyors over trucks could see conveyor waste systems deployed extensively in those mines where the deposit allows. These changes will fundamentally shift how strip mines are designed and operated with an accompanying continual improvement in mine safety and reduction in cost while, as always, improving underground operations will continue to offer a viable alternative.

REFERENCE

Bucyrus. 2008. Walking draglines: The range. www.bucyrus .com/media/23591/draglines%20trifold%200105.pdf. Accessed May 2010.

Strip Mining

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