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pter focuses on environmental issues specific to open pit mining one of the largest environmental impacts associated with open pit mining stems from the physical change of the landform as a result of the vast quantities of material moved spitz and trudinger 2008 the physical change of the landscape encompasses both the excavation itself and the disposal sites for overburden and tailings the excavation will have a large impact on the visual amenity of the landscape drainage patterns and groundwater levels furthermore the higher stripping ratios and generally lower ore grades mean that surface mines vastly exceed underground mines in the overburden volume that is generated both from ore processing and overburden as a result a larger surface footprint is taken up by the overburden embankments and tailings impoundment generated by open pit mining associated with the larger volume is not only a larger aesthetic impact on the landscape but also a higher potential risk of spillage of toxins into the environment additionally the changes in topography make the site more susceptible to erosion concurrent or postmining revegetation and placement of geotextiles on erosion prone surfaces can provide sufficient protection against both wind and water erosion chemical contaminants increased turbidity changes in flow patterns and higher susceptibility to flash floods are the main effects of open pit mining on the surface water regime spitz and trudinger 2008 chemical contamination and suspended solids can generally be removed in treatment plants channeling or rerouting streams away from vulnerable areas can be done to prevent erosion and mobilization of contaminants lowering of the groundwater level in an area due to mining may result in vegetation losses and consequent changes in fauna ground settlement and lower flow rates from spring fed surface water most detrimental effects of groundwater level control can only be effectively mitigated after mine closure effects of blasting include excessive vibration and air overpressure as well as dust fumes and possibly flyrock the maximum allowed peak particle velocity ppv depends on the vicinity of populated areas and national or regional regulations the u s bureau of mines found that cosmetic damage to houses can start at a ppv of 12 mm s at a frequency of 10 hz the onset of damage is dependent on both vibration frequency and construction quality of a building but in general it can be said that a higher frequency needs a higher ppv to be damaging siskind et al 1989 flyrock should be avoided altogether correct blast design should minimize environmental effects from blasting and good communication with local residents can reduce perception of blasting by the public in populated areas dust noise and road traffic effects are more pronounced than at remote mine sites noise causes disturbance of wildlife and annoyances both with residents and operators the emphasis on noi

se management should be on reducing noise and limiting exposure time department of consumer and employment protection of the government of western australia 2005 in this context hearing protection is regarded as an interim noise protection measure unless other measures are demonstrably impractical cabs on modern equipment are designed to reduce noise exposure and as such play an important role in limiting exposure time lower perception of noise by the public can only be achieved by taking noise reduction into account during production planning control of noise at the source and noise barriers typical exposure limits imposed by legislative bodies range from 80 to 90 db a for average exposure levels and peak exposure levels from 135 to 140 db a niosh 1998 depending on the particle size shape and chemical composition dust can cause physical or chemical contamination of equipment and soils as well as respiratory and dermatological problems reduced visibility and coating of vegetation department of resources energy and tourism australia 2008 the extent of the impacts of dust is highly dependent on climatic conditions and dust composition dust is generated by drilling and blasting excavation haulage dumping and processing of material or it can emanate from poorly vegetated or bare areas like tailings impoundments in combination with wind dust generated from drilling can be suppressed by adding water to the bailing air from the drill hole and by employing drill deck shrouds that envelop the drill stem haul roads are the most significant contributor of dust in surface mine operations emitting between 78 and 97 of all dust as discussed earlier in the section on haul roads dust suppression measures include spraying with water or chemical dust suppressants and compacting or changing wearing course material if loading hauling and dumping causes significant dust release wetting the material before excavation can be considered although this has detrimental effects on the efficiency of excavation and haulage water sprays around other sources such as stockpiles and the mill can be used to prevent fugitive dust from these areas in australia maximum exposure levels for pm10 dust with a median diameter 10 m lies at 50 g m3 on a 24 hour average department of resources energy and tourism australia 2008 toxicity of the particulate matter is dependent largely on silica content and exposure levels should be adjusted accordingly an increase in mining related surface traffic on public roads may cause congestion damaged or polluted roads dust noise and unsafe situations where unacceptable situations occur other forms of haulage careful route planning and mitigation of effects can relieve pressure on the road network but often the problems cannot be solved completely mine rehabilitation and closure the main objective of mine closure is to ensure physical and chemical stability as well as restoration of the

ecosystem in all areas disturbed by the mining operation spitz and trudinger 2008 before closure of a mine ore robbing or scavenging can be done this practice involves extraction of residual ore tied up in haul roads and other areas that did not allow extraction of the ore for operational reasons one example of ore scavenging was done at palabora south africa to supplement underground ore production it is important to realize that ore scavenging may reduce factors of safety for slopes to dangerous levels and should only be done after careful analysis of the probability extent and consequences of slope failure furthermore it may inhibit access into the pit should this prove necessary in the future mine closure aspects specific to open pit mining mostly arise from the large surface impact and large volumes moved spitz and trudinger 2008 careful planning during the development and operational stage and taking into account postmining use of an operation can significantly reduce postmining site disturbances and costs generally it is attempted to put the postmining landscape to an equal or better use compared to the premining landscape stabilization and erosion protection of steep slopes and unstable faces are crucial for ensuring physical and chemical stability this usually involves recontouring to stabilize pit faces and embankments to approximate original contours of the landscape followed by establishment of vegetation or liner placement for erosion protection flooding of the excavation allows it to be more successful at blending in with the landscape and natural ecosystem it prevents oxidation of sulfides and subsequent potential acid mine drainage and it aids in the prevention of reestablishment of drainage patterns backfilling is another option for rehabilitation of an excavation but may be a very costly option that is generally only considered when there is a depleted open pit mine nearby when there is an active excavation or other supply of large volume of material in close proximity to the pit to be backfilled or when backfilling is explicitly demanded by environmental regulations maintaining chemical stability of overburden and tailings typically involves the prevention of acid mine drainage and subsequent metal leaching as well as immobilization of any other toxic chemicals left over from the mining process spitz and trudinger 2008 flooding or sealing any potentially acidgenerating material prevents oxidation of sulfides and the resulting acidification of water if mine drainage is expected to contain any contaminants a treatment plant should be erected with sufficient capacity to treat all effluent water and possible storm surges open pit planning and design is a decision making process that leads to a realistic and actionable plan to profitably harvest mineral resources planning can be carried out for a wide range of time frames from the very short e g the next shift of mining activities to the ve

ry long e g the profitable extraction over the complete life of the mine this chapter is focused on the long term and on the creation of long term value it is about maximizing the value of the mine to the company excellence in long term mine planning involves making good decisions in pit design mine sequencing production rate process method ore selection and mining method all these decisions need to be made simultaneously and in conditions of great uncertainty such as the following are the geotechnical conditions well understood will the pit walls survive future seismic and weather events what is the resource under the ground really like how much ore is available and how will it respond to the planned processing methods what will commodity price and costs be next year in 30 years welcome to the very exciting and the very uncertain world of open pit mine planning and design to excel in this field the planning team will need to know something about mathematics and engineering business and economics and risk and uncertainty it is impossible to cover all of these subjects completely in one chapter therefore this chapter has been written to equip the reader with a basic framework and a good set of analytical tools to allow one to have a fair degree of competence in open pit mine planning the chapter is organized as follows planning process this section describes an overall planning process for open pit mine planning and design principles this section outlines guiding principles for the open pit mine planning process adoption of these principles is a major assumption for the rest of the chapter for example net present value npv maximization is established as the primary objective and this is treated as an assumption in the rest of the chapter open pit planning techniques this section provides information on basic techniques used in open pit mine planning techniques are presented in their approximate order of application models and tools this section provides information about model structures optimization and workshop techniques planning process the process described is generic and can be adapted for many types of projects including open pit mine planning the process is summarized in table 10 2 1 and the details are provided in the following discussion the process is described as a sequence of steps for clarity in practice however it is not uncommon to cycle back to earlier steps for example when new information becomes available in step 2 preparation of inputs it may give rise to a change in the purpose and scope of the study step 1 the whole process may be applied once or iteratively to cover the range from a broad study of resource potential to a narrow optimization study intended to generate a detailed actionable plan the difference between the broad and narrow application is reflected in the study s purpose and scope general preparation any open pit mine planning pro

cess requires leadership management and resources with responsibilities shared by a number of people example responsibilities and role assignments are shown in table 10 2 2 stakeholder analysis is conducted to determine the purpose and scope of the process to generate inputs to the study and to define the structure of the outputs engagement should be considered with all stakeholders including internal stakeholders e g senior management planning team members users of the plans and resource providers and external stakeholders e g investors customers government agencies and local communities the steering committee defines the purpose and scope of the study subject to the outcomes of the stakeholder analysis examples of purposes and scopes are provided for different study types in tables 10 2 3 and 10 2 4 preparation of inputs the next five subsections provide guidance on the main categories of information that should be prepared for the open pit planning process much of the information will be uncertain and one of the principles described later is to treat uncertainty explicitly one of the ways to deal with uncertain data is to employ the method described in the workshop for building a model of highly uncertain data section near the end of this chapter resource information information concerning exploration and mining rights in and near the area of interest must be prepared if the scope of the mine planning project extends to possible acquisitions then information should be gathered about the mineral rights held by others at nearby properties the rights held by others for alternate resources on the property should also be considered for example it is not uncommon for coal rights and coal bed methane rights to be issued separately although these rights apply to different resources it is very difficult to harvest one resource without impacting operations designed to harvest the other separately held rights can lead to very significant operational constraints and should be considered early in the planning process block models should be prepared for the areas of interest the reader is referred to the resource models subsection later in this chapter for information about block models current operations the following information about the current state of the operation needs to be obtained or prepared a process flow sheet encompassing mining operations and shipping operations if applicable bottlenecks in the business the current bottleneck the component in the flow sheet that determines overall system output with information about reliability and throughput variance growth bottlenecks what is standing in the way of mine expansion significant drivers of recovery dilution loss time in process and cost a cost model detailing fixed and variable costs for each part of the process flow sheet project alternatives information about potential technology alternatives for the expansion of an

existing mine or the development of a new mine must be prepared it may helpful to conduct a brainstorming session to generate a broad range of alternatives for consideration the reader is referred to the workshop for framing section near the end of this chapter for further information for each alternative information must be prepared that includes performance expectations expected throughput rates and variances etc and cost data capital fixed variable and terminal costs whenever the term expected is used in this chapter it should be understood to mean qualitatively that is the most likely if a mathematical definition is preferred then it should be defined as the probability weighted mean of all possible values for legal environment and community issues information should be prepared about mining and environmental laws applicable to the area of interest that will potentially limit the approach to mining or processing or affect costs in relation to local communities information should be prepared to address the following the potential for the community to provide people services or infrastructure the expectations of the community to share in the benefits of the mine proceeding the community s concerns regarding safety noise dust and impacts on the environment any other factors that may lead to the community being negatively or positively impacted by the mine in evaluating matters relating to legal environment and community it may be relatively easy to determine the minimum requirement for compliance and a desire to minimize costs may be the motivation to pursue that outcome however planning only for minimum compliance may not be the most effective way to progress a project the government and community are significant stakeholders and their support for the project will be strongly influenced by the mine s ability to meet their needs and expectations accordingly it is important to evaluate other alternatives that seek to actively garner the support of government and communities not only the costs associated with different alternatives should be considered but also their potential to avoid or cause delays and impacts on social license prices commodity price is the single most important variable impacting critical mine planning decisions it has a direct bearing on the optimal rate of production the final pit size and the desired degree of operational flexibility that is built into the operation if it were possible to know future prices with certainty and precision that knowledge could be applied as follows near term prices up to about 10 years into the future have the greatest impact on the determination of the optimal production rate for the mine the price at the end of the mine s life has the greatest impact on the size of the final pit the variable nature of the price measured over 3 to 36 months has the greatest impact on the degree of flexibility designed into the operat

ion commodity price is however the most difficult variable to predict many different forecasts can rationally be made for future prices and they all have one thing in common they are all almost certainly wrong because of this a mine should not rely on only one price forecast it should use at least the following three 1 mid price forecast the expected price 2 high price forecast a somewhat optimistic forecast 3 low price forecast a somewhat pessimistic forecast price forecasts are generally sourced from either the mining company s economist or an external consultant or research company whatever the source it is important to recognize that forecasts do not normally model all aspects of future price behavior this fact becomes apparent when a comparison is made between a chart showing historic prices and a corresponding price forecast chart e g see figure 10 2 1 they are characteristically different a historic price chart will show the price as it happened subject to the effects of market discontinuities changes in economic conditions investment cycles balances and imbalances between supply and demand and the influence of market speculators none of these price drivers can be predicted with certainty and some cannot be predicted at all forecasts can only be made based on scenarios the following are some example coal price scenarios declining coal resources will constrain supply leading to a price increase ongoing mining technology improvements will reduce coal industry costs thereby reducing the coal price coal supply currently exceeds demand and when marginal producers close down prices will rise emerging clean energy technology will reduce the demand for coal leading to price decline economic growth in developing nations will increase the demand for coal thereby increasing price consequently most price forecasts are in effect straight line assumptions modified by a scenario examples of simple forecasts are provided in figure 10 2 1 in addition to the simple models for price forecasts planners are increasingly using stochastic models for price in an effort to better plan for the random nature of prices methods include random walks and mean reverting random walks the use of these approaches can lead to better decisions as to the degree of flexibility in an operation however the detailed discussion of these methods is beyond the scope of this chapter when obtaining price forecasts one should inquire as to the following what aspects of future price behavior long run stable price long run trend near term reversion short run random are modeled in the forecast is the forecast in real or nominal currency are there intentional biases i e intentionally conservative or optimistic in the forecast most planning should proceed with the mid price forecast but the high and low price forecasts can be used to test sensitivities and to develop options most planners are familiar wit

h sensitivity analysis the idea is to determine how much a project value changes with a change in a major input variable sensitivity analysis often proceeds by varying the input variable by an arbitrary amount e g 10 however the results of the sensitivity analysis would be a lot more useful if the cases tested were more representative of the uncertainty in the input variable the highand low price forecasts should serve this purpose apart from their use in sensitivity analysis high price and low price forecasts can also be used to develop options discussed later in this chapter costs all costs need to be rendered in a structure suitable for cash flow modeling in one or more optimization systems planners should not rely on only one set of cost forecasts but instead should use at least the following three 1 mid cost forecast the expected costs 2 low cost forecast a somewhat optimistic forecast 3 high cost forecast a somewhat pessimistic forecast some costs are market based and the concepts and models described for forecasting prices may equally be applied to the forecasting of market based costs examples of market based costs include oil shipping equipment and construction services assumption sets to facilitate the assessment of alternatives it is very useful to define and document a number of assumption sets each assumption set consists of price cost and geological and operational settings the first assumption set should be the reference assumptions set and should include price forecasts at expected values the mid price forecast and cost recovery throughput rates geological and geotechnical models including pit slopes at expected values secondary assumption sets may be built to represent different scenarios for future conditions or resource quality volumes the following are examples high price assumption set this is the same as the reference assumptions set except that the high price forecast is used along with any associated impacts on costs low price assumption set this is the same as the reference assumptions set except that the low price forecast is used along with any associated impacts on costs high resource assumption set this is the same as the reference assumptions set except that the high price case from the resource range is used low resource assumption set this is the same as the reference assumptions set except that the low price case from the resource range is used each of the above mentioned assumption sets should have a reasonable 10 probability of being the most representative of current or future facts as analysis continues more assumption sets may be created for example in testing a broad range of production rates the impact of rate on mining width dilution and other factors should be carefully considered and appropriate assumptions should be determined and documented in some cases market based cost movements are correlated to commodity price movements

if that is the case then the correlation should be reflected in the high and low price assumption sets by including corresponding changes to the correlated costs framing the framing process leads to a set of high order decisions and alternatives to be tested as part of the study the result of the framing exercise is a strategy table an example of which is shown in table 10 2 5 assessment the framing exercise gives rise to a set of hypotheses possible best planning decisions to be tested each hypothesis should be subject to simulation and optimization to determine its economic value the initial assessment should be conducted using the reference assumptions set this assessment will reveal which set of decisions yields the highest value the structure of the conclusion is as follows if the reference assumptions set is true then the defined set of decisions are optimal example 1 using the reference assumptions a pit of 60 mt ore feeding a 5 mt processing plant is optimal assessment should continue with the application of other assumption sets to determine the optimal set of decisions for each the generalized structure of the conclusions is as follows if the named assumptions set is true then the defined set of decisions are optimal example 2 if the high price assumptions are true then a pit of 75 mt ore feeding an 8 mt processing plant is optimal the final stage of assessment involves identifying a list of candidate decisions to carry forward and subjecting these decisions to risk assessment selection the plan is selected based on the assessment results the following is a recommended structure best case plan a theoretical best case plan with few constraints and no secondary objectives applied this is generated using the reference assumptions set for further information the reader is referred to the establishing a theoretical best case plan section later in this chapter optimized plan the plan based on decisions that are optimal for the reference assumption set with all constraints and secondary objectives applied the npv change associated with the application of constraints and secondary objectives should be documented risk management actions these are modifications to the optimized plan designed to manage risks identified in the risk management process the cost to npv resulting from the application of risk management actions should be documented for further information the reader is referred to the identifying and managing risk section later in this chapter option preservation actions these are modifications to the plan designed to respond to alternate assumption sets a simple model for the preservation of options can be applied final plan this is the optimized plan as modified by the risk management and option preservation actions principles the following principles are posited as a foundation for openpit mine planning employ evidence based decision making methods the primary

objective is the maximization of npv treat secondary objectives as constraints do not apply public reporting rules to the input of the planning process deal with uncertainty and risk explicitly rather than by practicing conservatism details of each principle are discussed in the following four sections evidence based decision making evidence based decision making methods should be employed as follows seek to avoid bias in data gathering and in decision making assess information for its probability of being correct or predictive account for gaps in information and understand how those gaps contribute to uncertainty avoid using rules of thumb rules of thumb are simple approximations or rough methods for estimating or measuring something they are usually handed around from person to person without sufficient explanation of the limitations and hidden assumptions underpinning them such approximations are only useful if all their limitations and hidden assumptions are understood apply defined decision making processes and analysis methods for example applying the processes and methods described in this chapter will help to avoid making poor decisions learn to recognize and avoid logical fallacies an example of a logical fallacy is the mine must be either open pit or underground this example is called a false dichotomy because in reality there are other possible choices e g a combination of open pit and underground mining and in situ leaching interested readers can find information about logical fallacies on the internet primary and secondary objectives the primary objective is the maximization of npv this is the one goal that must be achieved subject only to the constraints of the secondary objectives the pursuit of the primary objective is one of optimization and only the maximum possible value will be satisfactory the primary objective is a matter of choice for a company engaged in the planning activity and a company is free to choose any primary objective it likes however in mining npv maximization is a very common primary objective and with a desire to elevate its readability this chapter has been framed with the assumption that npv maximization is the primary objective secondary objectives should be treated as constraints in the mathematical framework for optimization it is only possible to have one objective to be minimized or maximized however in business it is not uncommon to nominate multiple objectives to profitably employ the mathematical frameworks for optimization in open pit mine planning one of the objectives must be nominated as the primary e g maximize npv and all other objectives treated as constraints constraints are specified by minima or maxima or both and act to limit or modify the pursuit of the primary objective examples of secondary objectives are shown in table 10 2 6 secondary objectives applied as constraints cannot enhance the pursuit of the primary

objective they can only inhibit it a sacrifice is made on the primary objective i e npv is reduced when a secondary objective is applied secondary objectives should only be applied sparingly and the resultant sacrifice of npv should always be reported separation of public reporting from planning public reporting rules should not be applied to the inputs of the planning process this principle is particularly important when considering publicly reported mineral resources and ore reserves rules for public reporting are framed to provide a basis for consistency in reporting across all reporting entities and act as a filter on available information for example the joint ore reserves committee jorc code requires the competent person to make a preliminary judgment as to the technical and economic factors likely to influence the prospect of economic extraction and to apply those factors to ensure that the reported mineral resources have a reasonable prospect of economic extraction jorc 2004 the jorc code applies in australia but provisions are similar in many other jurisdictions the competent person s preliminary judgment as to economic factors acts as a filter to the full range of data regarding in situ mineralization it is much better for mine planners to be exposed to the full range of data regarding in situ mineralization including inferred resources and potential mineralization beyond inferred resources the planner can then apply more comprehensive economic criteria and analysis to determine the subset of in situ mineralization that is economically extractable when the planning process is complete the rules and conventions of public reporting should be applied to generate accurate public reports explicit treatment of uncertainty and risk uncertainty and risk should be dealt with explicitly rather than through the practice of conservatism it is common for mine planners to practice conservatism as a means of dealing with uncertainty or reducing risk two examples are 1 deliberately overestimating costs to compensate for the potential for costs to be accidentally underestimated and 2 deliberately underestimating the grades to compensate the potential for accidentally overestimating them acknowledge that there is uncertainty associated with most inputs to the mine planning process understand the inadequacy of single estimates of uncertain inputs even when a single estimate is an expected value its exclusive use can lead to poor planning decisions because it does not represent the probabilistic range an uncertain input is best represented as a continuous or discrete probability distribution and the full range of potential consequences of that distribution must be understood risk management should be applied when the full range of consequences of input uncertainty is understood open pit planning techniques basic technique for generating life of mine plans table 10 2 7 shows a recommended series of steps to

achieve the technical design of an open pit mine and its long term schedule the steps are iterative and at various points it is necessary to cycle through the steps in the pursuit of a good overall plan pit parameterization methods applying pit optimization to a single resource model and a set of assumptions will produce one pit outline the outline that maximizes the net cash flows undiscounted for a mine and often called the cash pit however it is useful to have a number of pit outlines available for analysis to be used as candidates for final pit designs and pushback designs a technique pioneered by matheron 1975 referred to as pit parameterization harnesses the pit optimization process to deliver a set of nesting pit shells that serve the purpose well several pit parameterization techniques are now in use each with varying advantages and disadvantages the following techniques are described in this section revenue factor cost factor vertical bounding and mining direction revenue factor in the revenue factor technique a range of factors is applied to the block revenue to alter its value in consecutive optimization runs the application of a lower revenue factor leads to ore becoming less valuable or being reclassified as waste and this in turn leads to a smaller pit a typical range of revenue factors is 0 10 through 1 00 in steps of 0 01 by running pit optimization repeatedly with incrementally different revenue factors a range of pits from large to small is obtained the pits so produced are often referred to as shells and are numbered 1 n from smallest to largest the shells do not overlap and are good candidates from which to choose pushback designs cost factor similar effects can be obtained by applying factors to either mining or processing costs or by direct manipulation of cutoffs a typical range of cost factors is 1 00 through 20 00 in 100 equal steps with higher cost factors yielding smaller pits the main advantage of this class of pit parameterization is that it usually gives rise to a large range of shells that if used as the basis of pushback designs can lead to high npv the main disadvantage is that the shells for vertically oriented ore bodies are often impractical with shells forming as concentric rings sharing no common wall vertical bounding the vertical bounding technique starts with all benches in the model being sterilized with the exception of the highest bench the pit optimizer is run only on the top bench and this may or may not lead to the creation of a pit depending on whether there is any ore in the top bench the process proceeds by progressively unsterilizing benches in the model and rerunning the pit optimizer the result is a set of nested pits and the advantage of this approach is that the shells that it produces can often have a better mining width at the base of pushbacks compared to the revenue factor or cost factor approaches the main disadvantage is that this approa

ch may provide too little or too much working space depending on the nature of the ore body and there is not much that can be done to make the optimizer produce a better outcome mining direction the mining direction technique is similar to the verticalbounding technique but the progressive unsterilization is done in a direction chosen by the planner in many circumstances the nested pit shells produced by the technique are more practical than those produced by other techniques as there is a propensity for shells to have a common wall the technique is implemented though user defined formulas in the pit optimization software acting to progressively unsterilize blocks according to their distance from a defined point or plane in the model the process is normally run several times using different origin points or planes the results can then be compared and evaluated based on their npv and practicality the main advantage of this technique is that it can quickly generate a range of alternative sets of pushbacks that are more practical to mine establishing a theoretical best case plan when applying optimization tools to open pit mine planning it is useful to start by generating plans that have as few constraints as possible this theoretical best case plan can be prepared as follows 1 generate multiple pit shells using the revenue factor method and use all of them as pushbacks this may lead to an unusually high number of pushbacks but that is of no concern when creating this theoretical best case plan 2 apply a simple schedule that limits prestripping e g one that mines out each pushback in turn or apply schedule optimization 3 only apply one production constraint at a time i e only one of mining process or selling limits and experiment with that limit over a wide range until npv is maximized 4 do not apply sinking rate limits or mining width constraints 5 ignore all secondary objectives the plan will almost certainly be impractical but it provides a very useful internal benchmark for the project the benchmark can be used to evaluate the npv impact of imposing secondary objectives and of the constraints applied to make the plan more practical when a practical plan can be created with an npv close to the theoretical best case plan then it is known to be a good plan practical and with a high value and there is little to be gained by further refining the pit designs and schedules determining the optimal production rate the optimal production rate is the rate at which npv for the project is maximized the simplest method to determine the optimal production rate is to calculate the npv for a range of different alternatives the testing is done by scheduling the mine for a range of different assumption sets with each assumption set representing a different production rate alternative with its associated capital variable costs fixed costs and selectivity pit geometry and infrastructure requirements as the rate

increases there are several impacts on operations and efficiencies in the pit and these should be reflected appropriately in the assumption sets the following are examples congestion will increase leading to longer cycle times unless additional mining width is provided larger equipment may be required leading to an increase in the size of the selective mining unit and an increase in dilution and or mining loss demands on ramps will increase leading to a need for wider or additional ramps the expenditure of capital unlocks productive capacity if the level of capital spending is increased it causes some upward and some downward impacts on value as follows the capital spending reduces the npv directly the capital spending increases production throughput which in turn increases net positive cash flows in each period and increases the npv the mine life is reduced thereby reducing the number of periods in which positive npv can occur and bringing forward terminal costs the fixed costs per period associated with maintaining the level of production may go up the variable costs associated with processing each unit of production may go down the results once graphed should be a convex curve such as the one shown in figure 10 2 2 the overall production rates for metals mines are ordinarily set by the comminution crushing and grinding function whereas for other types of mines the pace may be set by a different part of the value chain for example in an iron ore operation the pace is likely to be set by the capacity of the rail network or the port why are they different in each case the pace is set at the economic design bottleneck that is that part of the process chain that should be designed as the bottleneck of the operation the choice is based on examination of the capital costs and fixed costs of different parts of the process chain considering only capital costs for a moment the economic design bottleneck is that part of the process chain with the highest ratio of capital to product unit throughput for example if it costs 50 million to establish mining equipment and 100 million to establish process equipment to achieve a given level of product output then processing rather than mining is the design bottleneck in a metals mine the capital intensity of the comminution circuit normally outweighs the capital intensity of the mining operation or finished goods shipping therefore it is common for the pace in metals mines to be set in comminution in contrast the cost to install matching mining or finished goods shipping is relatively low so it is uncommon for these functions to constitute the economic design bottleneck for a metal mine in an iron ore operation the design bottleneck is likely to be in the rail or the port this is because the capital intensity for these parts of the value chain is relatively high fixed costs associated with maintaining a certain level of production also cont

ribute to the determination of the design bottleneck but it is beyond the scope of this chapter to discuss this in detail when the economic design bottleneck has been determined it should be used as the primary lever for testing increased or decreased production rates for the whole mine the design bottleneck constraint should only be applied in early stages of the analysis a wide range of production rates must be tested secondary constraints should only be introduced in later stages of the analysis and their impact on value and overall throughput should be carefully examined throughout the planning process focus should be maintained on the economic design bottleneck if a schedule emerges for consideration in which this constraint is underutilized in some periods then this is a strong indication that the schedule or plans are not as good as they should be determining processing cutoffs in practice mine planners rely on software to calculate or optimize the setting of processing cutoffs however it is important to understand the basic concepts and not merely place blind faith in the software to that end a simple explanation of some of the concepts is provided including the calculation of marginal cutoffs for simple situations and descriptions of some of the issues encountered when seeking to effect ore waste discrimination in more complex situations as part of the open pit planning process it is necessary to determine the treatment of run of mine material the alternatives typically include the following treat as ore and allocate to one of the available processing pathways treat as ore but defer processing by allocating the material to a grade stockpile the material would later be reclaimed for processing treat as waste and allocate to a waste dump the traditional method is to differentiate material by its product grade and the application of cutoffs in a simple case material with a grade above a defined product grade cutoff will be processed and material below the cutoff grade will be treated as waste the cutoff grade is calculated such that material will be processed only if processing will lead to a net increase in value if the objective is to maximize net cash flows the cutoff is calculated as the grade at which the net cash flow from processing is the same as the net cash flow from treating the material as waste the mathematics for this is shown in equations 10 2 1 through 10 2 6 the conventional term for this type of cutoff is marginal cutoff the simplest way to think about this is to consider material as it emerges from the pit at the pit rim with all mining costs paid the marginal cutoff is set at the grade at which the cost of processing exactly equals the revenue generated by the recovered product the following are some important observations as to the marginal cutoff shown in equation 10 2 6 the cutoff is independent of the mining cost the cutoff is independent of the grade distribution in

the ore body the derivation of equation 10 2 6 is easy to understand and this method can be applied in very simple cases if the simple objective maximization of net cash flows is acceptable however the mathematics becomes more complicated when multiple products or multiple processes are available this is because of the following multiple product mines require the application of multipleproduct grade cutoffs one cutoff per product figure 10 2 3 shows graphically how two product grade cutoffs are applied multiple process mines require the application of cutoffs for each process one cutoff per product per period per process figure 10 2 4 shows how multiple process cutoffs are applied when the objective is changed from maximization of net cash flows to the maximization of npv the calculations become intractable without the use of optimization software this is because decisions made in one period affect the discounted value of material processed in subsequent periods instead of calculating just one grade cutoff per product and per process a separate set is required for each time period if for example a mine has two products three processing methods and operates for 20 years then 120 cutoffs must be calculated and the setting of cutoffs in each period affects the optimal setting of all others when one further complexity is added the whole concept of discriminating based on product grades alone becomes problematic when the processing recovery for one product e g gold is affected by the grade of another product e g copper the application of product grade cutoffs as shown in figure 10 2 3 becomes inaccurate it can lead to the misclassification of some ore as waste and some waste as ore the solution is to use a cutoff method based on cash flows rather than product grades this allows for the avoidance of the misclassification described above and it also significantly reduces the number of decision variables in optimization software this is usually referred to as the cash flow method and it is very easy to apply when the objective is to maximize undiscounted net cash flows each parcel of material is assigned to the processing path that will maximize the profit or minimize the loss for the parcel if the objective is to maximize npv then a related approach called the dc method can be used the dc method extends the cash flow method used for marginal cutoff calculation and allows for its application to npv maximization problems figures 10 2 5 and 10 2 6 illustrate the application of the dc method to two processes mill and heap respectively the illustrations are for two processes only but the approach is applicable to any number of processes allowing for topographical features mine planners often need to deal with topographical features such as existing infrastructure roads and rivers that are difficult or expensive to move two methods are described here to deal with such features the first metho

d is forcing the pit optimizer to avoid undermining them and the second method is allowing the pit optimizevr to undermine them only if there is economic justification to do so the first technique involves the application of extremely high mining costs to certain blocks in the model and is effective at stopping a pit optimizer undermining defined topographical features figure 10 2 7 shows the surface level blocks in a model the shaded blocks represent the area that cannot be mined because of the presence of a river black line assigning extremely high mining costs e g 10 million per metric ton to these blocks will make them uneconomic to mine no matter how much ore is under them the second technique involves the definition of a set of arcs to model a topographical feature and to associate with this set the cost of relocating the feature arcs are normally used to model the pit slope constraints in pit optimization but their application can be extended to modeling other things including topographical features that are either expensive or impossible to mine through figure 10 2 8 shows how arcs can be used to represent such a constraint the dotted line shows the approximate southern boundary of the river the river may be relocated further to the north at a cost but without relocation the dotted line boundary is the furthest north that a pit can extend the arcs shown as arrows in the diagram trace a path along the boundary and then one final arc completes the loop the effect of this loop of arcs is that if a pit outline were to include any one of the linked blocks then all blocks must be mined the cost of relocating the river is built into the cost of mining one of the blocks in this example the shaded block it does not matter which block carries the cost provided that the cost is only carried once the effect of this arrangement is that the pit optimizer will mine through the boundary only if doing so will maximize the value of the pit i e only if relocating the river is economically justified allowing for safety berms and ramps pit optimizers cannot model safety berms and ramps directly instead an average pit slope angle must be calculated for the purpose of pit optimization which takes into account the safety berms and the width and number of times a ramp is expected to intersect a pit wall when the pit optimization work is complete the safety berms and ramps should be inserted into the design a geotechnical engineer prescribes the inter ramp angle it is the maximum slope that can be safely sustained between ramps and the base or the rim of the pit the geotechnical engineer will prescribe the vertical frequency and width of safety berms which must be taken into account in the calculation of the inter ramp angle the average pit slope for a given sector of pit wall is based on the inter ramp angle but adds an allowance for ramps to calculate the average pit slope for a given section of pit wall the foll

owing information must be known the design width of the ramps the width must take into account the types of equipment that must use it the corresponding height and position of the safety windrow the need for drainage and utilities and the requirements for geotechnical stability the number of times the ramp is expected to cross this section of wall the height of the wall from the base of the pit to the pit rim a plan for a final pit in the life of a mine the plan for the final pit may change many times the changes occur as new information becomes available for example new drill results that extend the resource revised price and cost forecasts or improved technology information constantly changes and to the greatest extent possible plans should be refined and improved over time to take advantage of the new information the final pit only comes into existence when mining stops and up until that time its final form is uncertain consequently the use of the term final pit plan is discouraged it implies that both the pit and the plan are final the plan is rarely final so the term plan for a final pit is preferred having established the plan for the final pit as a changeable thing for practical purposes it must manifest at any given time as something concrete something that can be tested with various economic assumptions schedules and mining rates and adjusted for different mining widths and ramp designs at the beginning of a planning project a reasonable proxy for the final pit can be found in the optimum pit outlines determined in the pit optimization process the pits corresponding to revenue factors 0 75 through 1 5 provide a good spectrum of pit sizes for consideration in the early part of the planning process in examining this range of pits the existence of common walls or floors must be determined common walls and floors indicate aspects of the pit shape that are stable for a range of economic assumptions these stable areas can be useful in mine planning but before relying on them as design features one must try to understand why they are stable it could be that the geological model includes a boundary that limits the ore zone and leads to a common wall in a range of pits or that the geological model shows ore extending only to a certain depth thereby limiting the depth of the pit in these situations the reliability of the geological information should be tested as follows could it be that the boundary indicated in the geological model is not exactly where it is shown could there be unexpected folds or faults is the vertical limit of the modeled ore based on positive information e g drilling penetrated barren rock below the ore or the absence of it e g drilling has not been extended below this level this sort of examination may prompt further data gathering or reinterpretation of the geology in the meantime planning should proceed as follows 1 use the largest pit in the range to estab

lish a reasonable maximum footprint this footprint will help in the positioning of infrastructure waste dumps and leach pads positioning these things outside the maximum footprint will reduce the risk of incurring the capital outlay to relocate them in the future however this risk reduction comes at a cost the cost of hauling waste and ore over greater distances the final decision as to distancing should be made on the balance of costs of longer haulage and the benefit of reducing the risk of future relocation 2 use the revenue factor 1 0 pit as a proxy for the final pit for the purposes of determining the mining schedule and production rate 3 as testing of the mining schedule and production rate proceeds retest pits in the range of revenue factors 0 9 through 1 0 and adopt the one with the highest npv as the new proxy final pit 4 refine the pit shapes based on ramp designs and mining widths produce a range of pit sizes and test them to determine the pit that yields the highest npv pushbacks two mine sequences are illustrated in figures 10 2 10 and 10 2 11 the first is simple to achieve with excellent mining access and low complexity however the stripping ratio in the early part of the mine life is very high and then decreases over time in this sequence much of the cost of stripping is incurred early in the mine life and access to ore is delayed this sequence leads to poor npv and for that reason is often referred to as a worst case sequence worst for npv in figure 10 2 11 a mining sequence with four pushbacks is shown by mining each pushback in turn the prestripping is deferred and ore access is achieved earlier the stripping ratio starts out low and ends up high this sequence is more complicated than the first but it will yield a higher npv the sequence shown in figure 10 2 11 is similar to the bestcase schedule described previously it has a many pushbacks and a high npv the trade off is practicality specifically too many pushbacks can lead to excessive expense to maintain multiple working slopes vertical advance rates that are difficult or impossible to achieve operational problems associated with moving equipment around multiple working faces lack of adequate mining width and extra cost associated with reworking ramps if these problems are experienced with a design that has many pushbacks it is likely that the npv calculated for that particular case is unachievable on the other hand if all pushbacks were eliminated practicality could be assured but for many ore bodies there will be a significant loss in npv how should the need for pushbacks be determined if pushbacks are needed how many should be employed the simplest way to determine the need for pushbacks is to experiment with different numbers of them the two extremes have already been discussed no pushbacks and many pushbacks the value difference between these two schedules will provide an indication as to the value adding potential

of pushbacks if the value difference between a best case and worst case schedule is small then there is little or no advantage in employing pushbacks however if the value difference between a best case and worst case schedule is great employing some number of pushbacks will add significant value in a best case schedule many pushbacks are used often in the range of 50 to 100 this is not a practical number but it is easy to generate this many and it provides a useful benchmark in practice it is possible to achieve very nearly the same value as a best case schedule merely by sizing each pushback to contain about a year s production however the success of this approach is to some extent illusory if npv is being calculated on annual rests which is usually the case for long term planning then the calculation of npv will be no different for example for 12 annual pushbacks compared to 48 quarterly pushbacks the npv calculation is insensitive to anything that happens on a timescale shorter than 1 year annual pushbacks are a good starting point but may still be more than is required to achieve a good npv therefore the recommended approach is as follows 1 compare the value of a best case schedule and a worstcase schedule if the values are similar then there is little advantage in employing pushbacks focus instead on practicality and operational efficiency 2 if the value difference between best case many pushbacks and worst case no pushbacks is great then an examination of pushback alternatives is warranted start to simplify the best case schedule by choosing pushbacks that each represent about a year of production also estimate the maximum number of pushbacks that can fit given the mining width constraint this is where practical requirements and theoretical best case npv collide if pushbacks are too small to be mined then the reported npv is theoretical proceed with no more pushbacks than can be practically achieved 3 experiment with reduced numbers of pushbacks adjusting their size to achieve the most favorable schedule if a stripping hurdle occurs in a schedule it will often be associated with the conclusion of one pushback and the commencement of another try to increase slightly the size of the concluding pushback or try decreasing the size of the commencing pushback to overcome the stripping hurdle test the npv of each alternative and settle on the smallest number of pushbacks that will achieve npv that is close to the best case benchmark allowing for mining width when a pit optimizer generates a set of pit shells it takes no account of mining width as a result there is no guarantee that any pair of shells will obey mining width constraints to the extent that good mine planning requires a mine planner can take care to choose pit shells that exhibit most of the attributes required but inevitably some intervention will be required to adjust the shape of the pits some software packages have options t

o apply mining width constraints subject to a user defined set of parameters without access to such systems manual design techniques should be applied to adjust for mining width whenever an adjustment is made to a pushback in one bench it can affect the shape of the pushback in benches below or above it refer to figure 10 2 12 any such changes to pushbacks will almost certainly lead to a loss of npv allowing for blending in the field of open pit mine planning few would doubt the enormous impact that the lerchs grossman pit optimization method has had the method is highly efficient and it can be guaranteed to produce an optimal solution however it can only produce an optimal solution if each block in the model is assigned an independent value and it is not always possible to do that this can certainly be the case when run of mine material is blended to make a salable product a simple three block model is shown in figure 10 2 13 in this example low grade ore can be blended with high grade ore in a 1 1 ratio to make salable product assuming all blocks contain identical tonnages of material the high grade block should be mined but which of the two low grade blocks should be mined either one or the other of the low grade blocks should be mined but not both there are two optimal solutions to this problem and no matter how the economic model is built it is unlikely that the lerchs grossman method will be able to find either solution so how does one define the pit shape for blended products lerchs grossman pit optimization can still be used to contribute to the solution but it cannot be relied on to the same extent as for nonblended products the lerchs grossman method must be supplemented by a technique to prepare models for optimization and test for the optimality of solutions these will be discussed in reverse order to describe the test for optimality it is necessary to first establish a typical pricing model for blended products figure 10 2 14 shows a typical pricing arrangement based on the product grade there is a target grade and a price applicable for blended product with that grade above the target a premium is paid and below the target a penalty is applied if the grade gets too low the product is rejected altogether in addition to being affected by product grade blended product price may also be affected by the grade of deleterious elements figure 10 2 15 shows a typical structure for the premium or penalty for a deleterious element if the grade is too high the product is rejected altogether in most cases the best value for a mine will be achieved when it can consistently produce blended product that just meets or slightly exceeds the target grade requirements this is because the gradient on the price curve changes at the target grade the rate at which a premium is paid is less than the rate at which penalties are applied if the product is worse than the target grade the implication for pit desi

gn is that the average grades product grade and deleterious element grades of all ore mined must equal the grades of the on specification product that sets an important condition for optimality of the pit that the average grades of all material flagged as ore and included in the pit must be equal to or slightly better than the target grades analyzing the case for flexibility the beginning of this chapter has focused on the long term trends for prices not the short term fluctuations however short term fluctuations will surely occur and mines need to test the impact and assess the case for building flexibility into an operation to allow it to respond to fluctuations an example of flexibility in a metals mine is the ability to change the processing cutoff grade when prices are high it may be advantageous to increase the processing cutoff this increases the output of product allowing the mine to take full advantage of the high price to be able to make this change the mine must have the ability to move quickly to a higher mining rate because raising the processing cutoff will lead to a higher strip ratio and to keep the rate of ore constant the mining rate must be increased the benefit of flexibility in this case is the ability to take advantage of the higher prices the cost is the additional cost of having mining flexibility price is the most important economic driver for all mines and it is also quite variable other important economic drivers may exist that are just as uncertain as price and it is important to analyze the case for flexibility the following are other examples providing excess capacity on ramps allows mining rates to be increased quickly allowing some flexibility in the comminution functions allows short term recovery and throughput changes to be made in response to unexpected ore characteristics maintaining spare mining capacity and an inventory of exposed ore provides more short term options to manage head grade and throughput identifying and managing risks the inputs to mine planning have varying levels of uncertainty and uncertainty can translate to risk steps must be taken to reduce risks to an acceptable level to achieve a successful outcome it is necessary to manage all risk aspects closely including the potential for risk reward that is the acceptance of additional risk within an approved tolerability limit resulting in increased rewards each mining company needs to develop an organizational risk management process and culture that provides its management and shareholders with certainty and confidence in its planning process organizational management processes are dealt with in detail in other literature whichever approach is adopted it is important that this process accomplish the following identifies assesses and manages both strategic risks associated with the mine plan as well as tactical and other risks that the plan must manage supports the planning process through the

objective analysis of the contributing influencing uncertainties and their consequences designs implements and monitors effective risk controls that either reduce the probability of the risk occurring or the consequence impact of an event when it occurs applies risk management equally to risk reduction avoidance and to opportunity upside potential the result of the analysis will be changes to the mine plan that impact risk and npv determining options for preservation an option in finance is the right but not the obligation to execute a transaction at a later time an option can be bought and has a value because it exposes the owner to the possibility of the later transaction in mine planning there are many decisions that are conceptually comparable to an option in finance for example maintaining flexibility in a mining fleet has a cost but it provides the ability in the future to quickly increase mining rate allowing the processing cutoff to be increased thereby increasing monthly product output if this option is available it is one that could be exercised during periods of very high commodity prices without having an option such as this the product output could not be changed in response to changes in the market an option must be exercisable in the following situations a credible future price regime response to positive exploration results response to a positive research and development outcome other circumstances that are credible and that have nonzero probabilities of occurring if it is possible to estimate the probability that the future option will become available the value of the exercisable option vo is calculated using equation 10 2 10 vo pe voe cpo 10 2 10 where pe probability that a trigger event will occur triggers the option voe value discounted cash flows of the option if it were exercised cpo cost discounted cash flows incurred in preserving the option if it is not possible to estimate the probability of a trigger occurring then an alternative is to calculate the minimum probability pe min that is required for the option value to be positive vo 0 if pe min is sufficiently low then it is reasonable to conclude that the option value will be positive table 10 2 8 shows an example of a preserved option determining the open pit underground interface for some ore bodies mining is possible by either open pit or underground methods if the choice is between applying one method and the other then a decision as to which method to choose can be made by comparing the best open pit plan with the best underground plan in the case that a combination of the two methods may be employed then a decision must be made as to where one method ends and the other begins when both open pit and underground mining methods are used a great many problems can arise these include the complexity of sequencing the commissioning and operation of the mines and the costs and risks of managing the

geotechnical and safety consequences of digging a pit near an underground mine the detailed analysis of these issues is beyond the scope of this chapter what follows is a description of a simple method to determine an interface between the two methods with some parts of the ore body being taken by the pit while others are left for underground mining the interface between the two methods will occur at a depth or location at which the open pit method becomes less economic that the underground method the determination of this interface can be achieved by applying a pit optimizer to a specially modified block model the modifications enable the calculation of the value of each block as the difference between its open pit value and its underground value this difference represents the net benefit of mining a block by openpit method instead of by underground method the following is an important logical assumption that underpins this analysis if a block could be mined by underground method and it is not mined by open pit method then it will be mined by underground method a block can be considered that has value x if mined by open pit method and a value y if mined by underground method the value y will be achieved if the block is not included in the pit because it will instead be mined by underground method the value x will be achieved if the block is included in the pit outline but the improvement in value is not x it is only x y this is the value that should be used for the purposes of pit optimization the difference in value between open pit and underground the calculation of the underground value y relies on the application of an underground planning process beyond the scope of this chapter to describe instead here are some guidelines as to how the underground value of blocks should be considered in the open pit mine planning process the calculation is described in principle only as the application details of this method depend on the modeling tools included with the pit optimization software being used the minimum requirements are described in table 10 2 9 the underground value y is equal to the aboveground value less the mining and haulage costs now that the underground value of each block is known it should be deducted from the value of each block in the model prior to pit optimization pit optimization proceeds with block values equal to the open pit value less the underground value x y the result will in most cases be a pit that is smaller than it would have been if the block values had not been adjusted it is possible for the pit to stay the same but it is not common that part of the underground mine plan that was not mined out by the open pit mine remains as an underground mine development costs are not included in the calculation of the underground block value these costs are important but there is no satisfactory method for apportioning the development costs to stope blocks in the model the alterna

tive is to deal with the development costs outside the pit optimization process if some blocks are mined in the open pit that could have been mined by underground method then the initial underground mine plan assumptions are no longer valid this is particularly true for development that is justified by its ability to gain access to several stope blocks if some of those stope blocks are removed from the underground mine plan then the justification for the underground development must be reassessed this reassessment may lead to further changes to the underground mine plan specifically the removal of any development that can no longer be justified and consequently the removal of any remaining stope blocks that rely on that development if this is the case then the whole process of determining the open pit underground interface should be repeated but now based on the new smaller underground mine plan another consideration is that the pit optimizer will take no account of the separation between open pit and underground mining without proper separation either the open pit or the underground or both mines will become unstable and unsafe a geotechnical engineer should be consulted and adjustments made to the plans to ensure that both plans can be safely executed anticipating and allowing for mine closure mine rehabilitation and closure are described in detail in other chapters of this handbook in this chapter the treatment of closure is limited to the cash flow effects and their impact on mine planning decisions table 10 2 10 shows different types of rehabilitation costs and their recommended treatment the treatment of costs shown in table 10 2 10 must be simplified for the purposes of pit optimization because in that process only variable costs can be modeled fixed costs e g interest associated with a rehabilitation bond should be treated as implicit fixed costs terminal costs should be treated in the same way as capital models and tools introduction to optimization optimization is a term that can be used in a general sense to mean a process through which an outcome is made as good as it possibly can be through the adjustment of inputs structures or methods the term also has a more precise mathematical definition which means to find the optimal value of a function often subject to constraints optimal value in mathematical terms means one of the following minimal value the lowest possible value to optimize total cost is to minimize it maximal value the highest possible value to optimize total profit is to maximize it a model for optimization includes the following objective this is the item to be optimized objective function the objective function is a mathematical expression or some other kind of mathematical model that calculates the thing that is to be optimized subject to the variables described next input data costs prices and efficiencies are all examples of input data on which th

e calculation of the objective value depends decision variables to optimize the objective an optimization process must find the appropriate settings for one or more decision variables an optimization problem may have one or many decision variables constraints the settings of the decision variables are subject to constraints both on the decision variables themselves and on functions of the decision variables optimization method this is a method for determining the appropriate settings for the decision variable such that the objective is optimized mine planning optimization tables 10 2 11 through 10 2 13 show basic descriptions of the most commonly applied optimization processes in open pit mine planning these optimization methods are available in a variety of commercial software packages costs mining processing transportation time in a bottleneck process assignment to an ore or waste stream pit slopes examples of significant drivers are provided in table 10 2 14 in practice the data available for inclusion in the block model will be less than ideal with data collection and modeling constrained by budget time or technology the gap between the information that is desired and the reality of available data should be examined at each phase of planning and recommendations made for exploration and geology so that sufficient modeling data is available for subsequent planning phases stockpile models the following definitions relate to the purpose of the stockpile and are represented as three different types although any given stockpile is usually of a single type it is certainly possible for a single stockpile to serve more than one purpose grade stockpile a grade stockpile is a stockpile maintained for the purpose of deferring the processing of lowgrade material until later in the mine life this is done principally to increase project npv grade stockpiles are large on the scale of years of production deposition and reclamation occur over many years grade stockpiles are usually associated with precious metals and base metals mines they are less commonly associated with bulk products e g coal iron ore and bauxite or industrial minerals e g limestone mineral sands and phosphates blending stockpile a blending stockpile is a stockpile maintained for the purpose of storing material with particular grade characteristics until such time as it can be blended with other material either from stockpiles or run of mine material so as to achieve a desired blended characteristic blending stockpiles are common for many bulk products and industrial minerals they may also be used to facilitate the feed to an extractive process for precious or base metals to improve processing recovery or throughput or to reduce costs buffer stockpiles buffer stockpiles are designed to deal with short term mismatches between the output of one process and the input to another they are relatively small representing

a few hours or days of production a common location for a buffer stockpile is on the front end of a comminution circuit to ensure that the process continues without being affected by fluctuations in deliveries of ore from the mine cost models for optimization to perform the optimization the impact of decision variables on revenues and costs must be incorporated into the model revenues are positive cash flows and are associated with product output costs are negative cash flows and must be associated with activities in such a way as a change in decisions made in the optimization model will be reflected by an appropriate change in costs the following basic modeling definitions are required to conduct the cost modeling effectively 1 capital costs these are costs that are incurred once to build or unlock productive capacity for example the cost to build an additional processing plant is a capital cost associated with the productive capacity of the new plant 2 variable costs these are costs that vary in proportion to activities such as mining processing and shipping finished product there is a generally accepted set of categories used in open pit mine planning for variable costs and these are described in table 10 2 16 3 fixed costs these are costs that are incurred regularly periodically and are associated with maintaining a level of productive capacity fixed costs are sometimes called time costs for example the cost to maintain a mining camp for 500 employees and their families is a fixed cost it does not increase or decrease with the rate of mining processing or product produced on a monthly basis and it only changes with the number of employees 4 terminal costs these are costs that are incurred at the cessation of normal mining operations typically mine closure and rehabilitation costs these four cost definitions and the more detailed variable cost categories in table 10 2 16 are all that are required to build a complete cost model for mine plan optimization it is important to make a distinction between the definition for capital cost shown previously and a range of other costs that use capital in their title for example terms such as replacement capital and sustaining capital refer to expenditure to maintain the productive capacity of capital assets and should be treated as fixed costs or variable costs capital costs for equipment that has a productive life that is shorter than the expected mine life can also be treated as a variable cost for example in a long life mine the truck fleet may be replaced several times and at the end of the mine life the remaining fleet will be sold at a price that relates to the trucks remaining productive life in this case the capital can be prorated into the variable cost of mining if a contractor is used for mining then the capital costs remain in the contractor s accounts and the only costs that are relevant to the contractee are the ones charged to the contracte

e these are framed primarily as variable costs sometimes costs can be modeled as either variable costs or fixed costs and the best method may depend on whether analysis is being done over a very long term or a very short term as an example when the cost of labor in mining is considered as the rate of mining increases and decreases over the life of the mine labor can be increased and decreased to meet the needs over the long term labor is a variable cost over short periods for instance from one month to the next if the need for mining labor changes there is very little that can be done to increase or decrease the supply of labor and its consequent cost over short periods labor is a fixed cost as it is for the example just presented so it is for many other costs in long term modeling many costs behave as variable costs but these same costs may behave more like fixed costs over the short term a blend of these two behaviors is also possible the semivariable cost however rather than thinking of this as a separate category of cost it can be thought of as a cost in two parts one part is variable and the other is fixed alternate method for modeling costs for bottlenecks in the cost models for optimization section all costs are modeled as capital fixed variable or terminal costs a fixed cost varies only with time and a variable cost varies in proportion to activities such as mining processing and shipping of finished product one of these activities will constitute the bottleneck for the business and for the bottleneck instead of applying a variable cost model it is possible to apply a special kind of fixed cost model this is particularly useful when the throughput rate for the bottleneck depends on the type of material sent to it for example in a processing plant a series of parallel adjacent pits figure 10 3 3 shows a typical dragline pit overburden material from the current pit is placed in the previous adjacent pit from which product has been removed by auxiliary equipment pits are narrow and relatively long pit widths are most commonly 25 60 m 80 200 ft widths for rehandle operations tend be on the large side to reduce the percentage of rehandle width is influenced by the maneuverability of the product removal equipment depth of the overburden blasting method material characteristics dragline advance rate and dragline dump radius pit lengths vary greatly because of the influence of geology topography and artificial obstacles they are most commonly 1 000 2 000 m 3 000 6 000 ft although some operations have used pits as short as 300 m 1 000 ft or as long as 3 000 m 10 000 ft in shorter pits the sequencing of product removal and blasting becomes complicated and frequent ramp construction is required in longer pits power distribution systems become expensive and complex and dragline propel distances can be excessive the pad on which a dragline sits while it works must be clear of

hard spots and protruding rocks and must be relatively level graded to a slope of 2 to provide drainage yet avoid damage due to swing motor overheating and structural stresses modern draglines can propel up and down a 10 grade or across a 5 grade when they transition between grades it is important that they do so gradually always distributing the load evenly across the tub the dragline s circular base and shoes as a general rule the rate of grade change should be 3 per tub diameter for example for a tub 20 m in diameter the rate of grade change should be 3 per 20 m so a change from 0 to 9 should take at least 9 3 20 60 m additionally when the possibility exists of bridging the shoes the pad material should be sufficiently compacted to prevent supporting the shoes by the endpoints only draglines are designed to work in soft underfoot conditions and as such are designed with tub ground bearing pressures on the order of 1 2 1 4 kg cm2 17 20 psi during propel about 80 of the machine weight is transferred to the shoes and the remaining weight is carried by the tub edge this ratio can be changed by carrying the bucket or setting it on the ground reducing the tub edge load by setting the bucket on the ground reduces the probability of pulling a roll under the tub during propel in soft underfoot conditions rather than remove material from a continuously advancing face as a shovel does a dragline removes material from a specified length of the pit called a set or block the dragline swings approximately 90 and casts into a pile in the previous pit set lengths for larger machines are about 30 m 100 ft or about 16 steps for the dragline to remove the overburden from a simple set a dragline may use from two to four tub positions before retreating to start a new set commonly a dragline follows a pattern of digging positions to excavate a set figure 10 3 4 the first two rear positions are set back far enough to ensure that no material is too close to the fairlead to be reached in shallow pits these first two positions may suffice to reach the desired depth in deeper pits digging may soon reach the point where the drag ropes scour through the crest of the digging face in which case the dragline must move forward to clear the drag ropes thus from the rear positions the upper part or lift of the set is removed in the last two positions the dragline has moved forward to the edge of the digging face to reach down for the lower lift of the set the lateral positions in a pit are also significant from the positions along the highwall 1 and 3 at the bottom of figure 10 3 4 the dragline can fix the alignment and slope of the new highwall with the key cut this trench like cut is confined to a bottom width of only a single bucket as it works down such a cut allows maximum lateral control of the bucket with minimal lateral strain on the boom in addition if the entire dig path is not directly radial

to the dragline production and mechanical availability can suffer at the bottom of the lift in position 1 the dragline generally moves laterally to position 2 this plug position allows the dragline to spoil at maximum range this move is made before completion of the key to reduce delay caused by hoisting clear of the key before beginning the swing as excavation progresses the plug is removed in lifts comprised of a series of cuts to an equal depth about one half the bucket height made in a sweeping pattern the sweeps normally progress from the spoil side to the key so as to minimize any hoisting required before swinging the bucket to the spoil then the dragline moves forward into positions 3 and 4 to excavate the lower lift in much the same fashion in figure 10 3 3 the dragline has stepped back from the face for maintenance but has completed the first three positions of the set the top of the set has been removed from positions 1 and 2 and the key cut has been finished from position 3 the dragline is now ready to move over closer to the spoil and finish the plug from position 4 this description although typical should be considered general set lengths and digging positions will vary depending on operating conditions and machine capabilities dragline operating methods draglines can operate by means of several operating methods described in the following paragraphs simple side casting the standard dragline method is simple side casting used when the dragline has the required reach to move the overburden to its final place with typical angles of repose and pit widths the maximum overburden that can be handled by this method is a little less than half the effective radius discussed later in the dragline selection section of this chapter variations on simple side casting are common advance benching advance benching figure 10 3 5a is useful in areas of uneven terrain or in overburdens where a top layer of unconsolidated material overlays competent rock the set is split into an upper and lower bench the lower bench is removed conventionally the upper bench is removed by chop cutting which typically means digging above the working level but also can mean engaging the bucket at the dump radius either way the bucket is at least partly pulled down a face rather than up or across it the bucket is usually held in a dump position teeth down then lowered onto the face and dragged in chop cutting is sometimes used instead of a key cut in spoil side operations to clean the highwall however chop cutting is hard on the rigging ropes and bucket and can increase downtime and repair costs and decrease productivity productivity is further decreased by the lower fill factor and increased drag to fill time advance benching generally requires a longer swing angle as well although efficiency varies typically a reduction of 10 20 of the conventional rate should be used for initial estimates when practical chop cutting

should not be done above the height of the fairlead or else production and maintenance can be significantly impacted if the material in the advance bench is extremely unconsolidated it is sometimes convenient to build a buckwall visible in figure 10 3 5a out of dry competent material removed from another area of the set and placed as a retaining wall at the toe of the spoil the unconsolidated material is then placed and contained behind the buckwall a buckwall can be used to help stabilize any spoil pile extended benching to extend operations in deeper overburdens the alternative methods of extended benching and spoil side benching discussed in the next subsection can be used to remove material to a depth of about twice that achievable by simple side casting however these methods come at the price of increased rehandle slower cycles and more complex planning although the impact on rehandle and cycle time can be reduced by use of auxiliary equipment these methods can also be used on a temporary or localized basis around ramps spoil or highwall slumps at high spots in the overburden or inside curves in extended benching figure 10 3 5b the dragline places the driest most competent material from the set against the old highwall enough material is placed so that after leveling by dozers it forms a bench the dragline then moves out onto the bench in a position closer to the spoil as excavation progresses the bench is removed this method can be used in two seam operations as well a disadvantage of extended benching is that the swing angle is long lowering total production when calculating production requirements it is important to remember that the rehandle material in the extended bench is loose material and therefore has a different bucket factor for long term applications extended benching is frequently combined with cast blasting and push dozing figure 10 3 6a both of which are effective for moving material short distances downhill blasting lowers the bench level decreasing rehandle it also moves some material to its final place increasing production the cast blasting profile is then leveled by dozers to form the extended bench bench height and width should be designed to take maximum advantage of dragline hoisting however although high hoisting can lengthen cycles the benefit of a lower bench height with its lower rehandle is usually the determining factor in setting bench height in multiple seam operations the bench height may be predetermined by the upper seam figure 10 3 6b two positioning issues often arise in extended benching cleaning the coal toe and cutting the key for cleaning the coal toe because the extended bench covers the coal toe it is clearly advantageous especially with thicker seams to position the dragline just outside the edge of the coal which is the best position from which to clean the spoil toe away from the coal toe and thus minimize rib loss however doing so

may require pushing the extended bench out a little farther than is necessary to meet dump requirements for cutting the key a lower bench without a setback prevents the dragline from being positioned directly over the key the dragline can be positioned no closer to the new highwall than the rear end clearance radius to keep the highwall clean and well defined a presplit blast is commonly used the dragline can be positioned out on the extended bench at a distance from the highwall equivalent to the dump radius and then the key can be chop cut alternatively the key can be developed byauxiliary equipment typically dozers but sometimes backhoes or smaller draglines ingle and humphrey 2004 in the latter case it is critical to consider in advance the finer blasting fragmentation required for auxiliary equipment spoil side benching in spoil side benching figures 10 3 7 and 10 3 1 also called pull back overburden is removed in two independent passes this method is common in two seam operations and is virtually required in three seam operations on the first pass from the highwall side material is moved by standard side casting spoil from that pass is allowed to ride up the highwall and then the peak is leveled to form a pad for the second pass the first pass is generally completed for an extended length of pit whereupon the dragline bridges across to the spoil bench on the second pass the spoil bench can be removed in either direction depending on pit sequencing spoil bench development and cable layout on the spoil side the dragline is positioned so that the key can be chop cut with the boom perpendicular to the highwall the design height of the spoil bench which also determines its width is based on the reach requirement to chop cut the key on the spoil side pass the dig depth and if the coal toe must be cleaned the dump height limitations of the dragline and tub position if the first pass does not generate enough material to achieve the necessary spoil bench height material can be removed from the spoil side position and placed one or two sets behind the dragline much like an extended bench it is best to develop the spoil bench with auxiliary equipment however the amount of material to be moved may justify assistance from the dragline and since the bench is developed in advance the dragline is chopping in the direction of travel it is better to develop a finish grade for only the road width of the dragline along the spoil edge leaving the outer edge of the bench for the auxiliary equipment to finish grade spoil side benching requires carefully managed cable moves and layouts particularly when raising the bench level from the spoil side with the cable on the bench and the dragline progressing toward the cable the bench must be raised in halves this requires moving the cable from side to side and sometimes also swinging over the cable swinging over the cable should be done only with a protective co

vering and carefully controlled operator technique spoil side benching enables operation of multiple draglines in a pit figure 10 3 1 in a fairly short pit this can afford a very high production rate however it is difficult to schedule multiple draglines primarily because of the complexity of matching the advance rate of draglines working on separate benches although production requirements can be proportioned short term variability in production rates invariably causes inefficiencies very few operations run tandem draglines in this manner for any length of time dragline production the amount of material moved by a dragline is determined by the following basic parameters bucket capacity how much material is put in the bucket cycle time how fast the bucket is cycled operating hours how many hours per year the dragline is kept digging bucket capacity bucket capacity was historically measured in accordance with sae standard j67 1998 this standard calls for a struck top and front face calculation and then subtracts 10 of the calculated volume to account for the slope of the front face it also uses estimating factors to assist with the calculations of the complex curves of a typical bucket however the final so called rated capacity in no way represents the behavior of material in the bucket rather it simply provides a uniform method for comparing bucket capacities for calculating production the rated capacity must be adjusted for material swell and fill characteristics the material in a bucket is loose so the rated capacity of a bucket can be thought of in terms of loose cubic meters lcm or loose cubic yards lcy material swell changes with handling and varies within a bucket or pile of material within a large pile it also varies with time additionally it can differ dramatically from the swell in a shovel dipper or truck body to account for swell and fill two factors have been introduced 1 the swell factor fs is affected most noticeably by fragmentation and material composition but also by bucket design 2 the fill factor ff although more complicated to determine generally has a larger impact on production variability and should be studied carefully fill is affected most noticeably by fragmentation and operator technique but also by bucket rigging configuration figure 10 3 8 and design fs and ff depend on material characteristics digging method rigging configuration and bucket design they do not remain constant if any of these variables change they are thus best determined from field data optimally by starting with bucket count and block volume and then for a particular material operating method and rigging setup calculating the bank volume moved per bucket fs for material in a dragline bucket is typically about 1 3 loose volume per bank volume ff is typically about 0 90 of the sae j67 rated capacity fs and ff can be combined into a single bucket factor typically about 0 70 mean

ing that a bucket rated at 100 m3 typically carries about 70 bcm the typical distribution of payloads is about 10 cycle time the digging cycle of a dragline is comprised of five main components 1 drag to fill 2 hoist and swing 3 dump 4 return swing and lower and 5 bucket spot the time required for each varies depending on a number of factors most notably dig depth hoist height and swing angle other variables include material characteristics dragline performance speeds and operator proficiency because of the diversity of these factors even machines of the same model and design have different cycle times typical designed cycle times for larger machines are in the range 50 60 seconds for a 90 swing with a low dump a typical cycle is dominated by components 2 and 4 about 70 of a typical cycle is required to get the bucket over to the spoil and back again component 2 of the cycle hoist and swing is actually three independent movements swinging hoisting and paying out drag each has a specific time requirement for almost any dump point one of these movements takes more time than the others therefore two movements are retarded intentionally so that the slower dependent movement has time to coincide at the dump point however drag pay speeds are so rapid that they are seldom the dependent movement it is convenient to think in terms of the curve that the bucket follows at maximum hoist and swing speeds this swing hoist coincidental curve represents the points in space that make maximum use of the time available for both functions cycles dumped below the curve are swing dependent cycles dumped above the curve are hoist dependent for example if a particular cycle is a long swing with a short hoist then it is below the curve and swing dependent hoist distance and swing angle can be controlled by pit and digging pattern design and performance speeds can be affected by operating technique thus it becomes obvious that a dragline operator should dump each bucket not at the peak of the spoil but rather near the swing hoist coincidental curve although this is not practical for every cycle the more coincidental cycles that occur the more efficient the operation hoist distance and swing angle are minimized by optimizing the bench height and the digging positions of the machine while minimizing the number of relocations which cause nonproductive propel time obviously machine positioning is limited by the key cut drag rope clearance tail clearance and reach requirements bucket speed during these independent movements is very dependent on operating technique hoist speed is a function of the load in the hoist rope which is directly related to bucket position because hoist acceleration is extremely quick and represents a small proportion of the hoist time acceleration time does not affect hoist time significantly to keep the bucket in the carry position tension must be maintained on the drag rope wh

ich increases the load in the hoist rope typically the hoist load is about 120 of the bucket and payload weight the closer the bucket is carried to the boom the greater the hoist load which ranges from about 110 to 140 the change in hoist speed is directly proportional to the change in hoist load so the further out the bucket is carried the faster it hoists however carrying the bucket too close to the dump radius generally causes material to slough off the front of the bucket reducing the fill factor in contrast swing time is very dependent on acceleration acceleration to full speed and deceleration to stop requires about 60 of the swing which is about 85 of the swing time for a swing angle of 90 acceleration and deceleration are relative to the rotational inertia which is a function of the mass times the square of the radius of the center of mass so it is greatly affected by bucket location on longer swings where hoist is not a factor keeping the bucket in tight as long as possible improves swing time of bucket capacity cycle time and operating hours inefficiency in cycle time is the most difficult to diagnose and improve operating hours ostensibly the easiest performance parameter to measure is time in reality there are more ways to categorize delays and define losses than there are mines around the world the ubiquitous hour is possibly the single most misleading term used in mining the basic goal is to reduce the number of hours we have to work with down to the number of hours actually worked clearly understood definitions are critical to production reporting and estimating figure 10 3 9 shows the relationship between calendar hours and operating hours although there are about 8 760 calendar hours in a year draglines typically operate for 6 000 7 000 hours per year for initial estimating it is adequate to combine availability and utilization values into a single operating efficiency factor table 10 3 1 typical availability and utilization values are each about 85 which provides an operating efficiency of about 72 of the scheduled hours availability and particularly utilization can be affected significantly by application the same dragline on a highwall pass will see a significant difference in operating efficiency on a spoil side pass where it is chop cutting and pad building typically due to increased propel time decreased dump rope life increased bucket maintenance and so on apart from the obvious minimization of maintenance and utilization delays for maximum production it is key to use the available hours efficiently most dragline operations can realize significant production gains by reducing cycles that do not contribute directly to ore production the most significant contributors to inefficient operation are rehandle nonproductive cycles and poor planning to address these contributors the following are important effective management of rehandle the use of auxiliary equi

pment and careful planning are the best tools to manage rehandle the cost of rehandle should be thought of in terms of incremental costs of ore production rather than the difference in cost per unit of production for dragline operation compared to that for auxiliary system operation in addition in certain applications a slight increase in rehandle e g changing a bench level or pushing out an extended bench can improve production due to the effect on cycle and propel time productive digging nonproductive cycles are often difficult to measure but can decrease production significantly there probably isn t an operation at work today that couldn t reduce the amount of time the dragline does pad preparation and cleanup work by better use of auxiliary equipment draglines should not waste time heeling the bucket to level the pad or pushing down the roll in addition efficient coordination of the ground crew during cable layout and pad preparation can minimize move times quality operation planning and coordination the amount of detailed planning required to operate a dragline most efficiently should not be underestimated particularly for nonroutine digging specific tub positions bench levels and material placement need engineered plans it is important to involve dragline crews in planning to ensure their understanding and consensus in addition it is a mistake to assume that everyone can easily create from a two dimensional drawing a threedimensional 3 d operation rather it is wise to use 3 d software or sandbox models to work through operations in advance especially complex operations such as ramp crossings in complex operations the use of playbooks with a diagram for every tub position to describe the digging and dumping points can be extremely valuable dragline selection an important concept to keep in mind when sizing or selecting a dragline is to select the dragline for the mine plan not the mine plan for the dragline draglines are engineered systems generally customized for an application even when purchased used they can be modified during reassembly to better fit an application the two major parameters used to select a dragline are dump radius and allowable load occasionally other parameters such as ground bearing pressure of the tub rearend swing clearance dump height and dig depth may also affect selection dump radius rd is the horizontal distance from a machine s center of rotation to the hoist rope when the bucket is vertically suspended part of this radius is consumed by the stand off distance so figure 10 3 10 which is the distance from the center of rotation to the crest of the old highwall the remaining dump radius is the effective radius re thus rd so re stand off distance varies depending on machine size operational history and overburden conditions in the absence of field data for planning purposes the minimum stand off distance is commonly considered to be 50 of the wi

dth of the dragline from the outside of the shoes or 75 of the tub diameter the allowable load sometimes called the maximum suspended load or rated suspended load is the maximum weight of bucket rigging and material for which the dragline is designed to provide optimum performance standard duty buckets including rigging weigh about 1 2 t metric tons per rated m3 2 000 lb per rated yd3 of capacity although they may vary from 1 1 to 1 4 t 1 800 to 2 300 lb depending on the application overburden densities are site specific but 1 8 t lcm 3 000 lb lcy is a common approximation the combined weight of a standard bucket and material load is then about 3 0 t per rated m3 5 000 lb per rated yd3 therefore an operation that requires a dragline with a 46 m3 60 yd3 bucket requires the dragline to have about a 138 t 300 000 lb allowable load allowable load is calculated for a 100 full bucket peak bucket load even though the average fill used for production calculations is less e g 90 for a given dragline model or more correctly a given frame size the allowable load can be varied by changing the dump radius basically a shorter reach means a larger bucket this design change is limited by the hoist load that the gearing and motors can handle maximum allowable load or conversely the longest boom that the frame can handle the general rule is the longer the reach the deeper the depth that can be handled without rehandle and the less rehandle however the best choice is almost always the largest bucket at a shorter reach despite the higher rehandle the larger bucket more than makes up for the additional rehandle with additional production by calculating bucket requirements using a standard range of production values 8 400 scheduled hours 60 second cycles 72 operating efficiency and a 70 bucket factor it can be calculated that a dragline produces about 250 000 bcm yr per m3 250 000 bcy yr per yd3 of rated bucket capacity this unit of production per unit of bucket capacity is a convenient general factor called the production factor or digging index humphrey 1990 and can be used to quickly estimate the annual capacity of a given bucket or conversely the bucket required for a given production the production factor is commonly used to compare differences in dragline operations either one application vs another or one dragline vs another with historical data production factors for specific applications can be developed over time and used for planning and forecasting purposes operating mines commonly have production factors of 200 000 300 000 depending on application cycle time and efficiency production factors are often calculated on an hourly basis with care given of course to which kind of hour is used the hourly production factor can be useful for measuring the efficiency of a specific operation within an application kennedy 1990 the application of production factor presumes that overb

urden production requirements are known and have been adjusted for any ore losses in the pit and plant and any rehandle expected for the scenario caution should be used with volumetric terminology some operations mostly outside of north america report production in terms of prime also called virgin or in situ volumes but label them bank volumes the prime volume is the actual volume of overburden above the coal that was uncovered and does not include rehandle in fact most mines typically experience 5 10 of additional operation rehandle for ramps bench fill and the like a production factor based on prime volume is thus smaller than can be expected in actual operation prime volume is adequate for comparing machines using the same method at the same mine however it is not a true measure of individual dragline productivity it is advisable to use the terminology total including rehandle bank cubic meters tbcm and prime not including rehandle bank cubic meters pbcm or for cubic yards tbcy and pbcy draglines as loaders although draglines are normally used to direct cast material they are also sometimes used as loaders in that case the normally imprecise dumping technique of the dragline must be altered to something as precise as is used when spotting the bucket in the dig most experienced operators have little difficulty adapting to point dumping and the method has been used to load hoppers trucks and barges generally it is most useful when the pit bottom is unsuitable for truck traffic and too deep for a hydraulic excavator for example it is common in florida phosphates for a dragline to mine the phosphate matrix and dump it into a slurry sump on the highwall the sump is about the size of a large mining truck body contrary to the initial mental image of this method the bucket does not swing over the target rather the swing is stopped with the bucket in the carry position and the drag is payed out to dump trucks are positioned facing away from the dragline so that the bucket enters over the tailgate the precision required to determine the truck position i e to spot the truck is no different than is required for any other loading tool however because the spot cannot be marked by the bucket as in shovel operation position markers may have to be provided as the dump target is generally on the highwall side it is easy to locate the target so that the dragline has a very short swing angle and a short cycle time bucket wheel excavator systems the bucket wheel excavator bwe figure 10 3 11 is one of the grand machines of the mining industry and traces its origins to drawings by leonardo da vinci the original concept from the late 1800s was technologically challenged by advances in the steam shovel had its practical beginnings in the early 1900s but had its first real mining applications in german lignite mines during world war i rasper 1975 for purposes of comparison the bwe system is a highca

pital cost low operating cost system that has limited flexibility and can operate through a limited range of applications with sensitivity to geologic variance bwes are highly customized and vary in design more than do any other mining machines to the extent that nearly every machine is not just unique but almost dissimilar the machines are very robust in design and consequently very long lived their most common mining applications are mining unconsolidated overburdens and lignite handling bulk materials such as stockpiles and load out facilities and heap leaching pad construction and removal three major types of bwe exist in mining operations 1 systems that direct feed into a shiftable conveyor system that connects to a series of other conveyors and a discharge system these machines can weigh up to about 12 500 t 14 000 st and cut material from a bank height of more than 50 m 160 ft at rates of more than 10 000 m3 h 13 000 yd3 h 2 systems with long discharge booms that direct place material into the spoil 3 compact systems generally two crawler machines used for bench heights of about 10 15 m 30 50 ft with production rates of 1 000 2 000 m3 h 1 300 2 600 yd3 h loader and hauler systems more material is moved by loaders and truck haulers than by all other excavation systems combined the deciding factors in the selection of this system are typically the qualitative characteristics of flexibility and the probability of achieving production and cost targets loaders and truck haulers excel in flexibility they are not dimensionally constrained by operating method and so are able to move in any direction for any distance they can thus work in constrained or irregular geology and terrain and can be added incrementally both of which make them virtually the only choice for use in very deep pit mines their flexibility also enables a mining operation to adapt quickly to changes in commodity prices geology and other influences that cause the original mine plan to change as it inevitably does because haulers always operate in parallel and in large systems loaders do so as well the impact of adverse performance on the part of any individual unit is minimized additionally the system can operate through a large range of geologies and climates this all provides for a dependable system with little unpredictable variation in efficiency production or cost for purposes of comparison the loader and hauler system is a low capital cost high operating cost system that is very flexible and can operate through a broad range of applications with low sensitivity to geologic variance loading tools loading tools are a specific class of excavator that depend on a separate independent haulage system the most common types of loading tool from small to large capacity are wheel loader figure 10 3 13 hydraulic shovel figure 10 3 14 and mining shovel figure 10 3 15 in mining they are used in conjunction with haulers most oft

en off highway mining trucks in the past 30 years and particularly the past 10 years the development of larger more reliable wheel loaders and hydraulic shovels has encroached into what was previously the exclusive territory of mining shovels the loading of large off highway trucks these trucks also called large mining trucks generally have capacities of 135 t 150 st the growth in wheel loaders and hydraulic shovels has lead to a new delineation in the loader market with wheel loaders predominant at the lower end hydraulic shovels in the middle range and mining shovels at the upper end of bucket capacity figure 10 3 16 selection because all loading tools perform the same basic function i e they load trucks the differences among them lie in other characteristics likewise due to design differences and resulting differences in capital costs and operating costs per unit of production evaluations of net present value can prove useless deciding factors are more likely the qualitative characteristics of capacity mobility flexibility life and support requirements capacity the range in capacity of the three types of loading tool is a differentiating characteristic for operation with large mining trucks their capacities are as follows wheel loader 27 45 t 30 50 st hydraulic shovel 27 81 t 30 90 st mining shovel 54 110 t 60 120 st these differences in capacity spread even further when compared on an annual production basis because of cycle time and operating hour differences generally when comparing machines of similar size the mining shovel cycles more times per year than does a hydraulic shovel which in turn cycles more times than does a wheel loader this is due primarily to differences in operating hours but also to a degree in cycle time production is most heavily influenced by the degree of utilization i e the extent to which a tool is kept in use when it is mechanically available it is also influenced by consistency and efficiency of application with problems arising when the pit layout is poor resulting in long swing angles excessive moves and other workarounds table 10 3 2 shows the influence of swing angle on cycle time and hence production level of a loading tool the optimal production factors for loading tools are approximately as follows the number in units of bcy per year per cubic yard is the same as the number in units of bcm per year per cubic meter production factors are discussed in the dragline selection section wheel loader 330 000 bcm yr per m3 of dipper capacity hydraulic shovel 350 000 bcm yr per m3 of dipper capacity mining shovel 400 000 bcm yr per m3 of dipper capacity mobility if mobility is critical to an operation the best choice is usually a wheel loader in operations having multiple faces that require frequent relocations or want a backup unit for multiple loaders the wheel loader is uniquely capable of rapid relocation however recent d

evelopments in larger low boy or float transporters has extended the capability of medium to large hydraulic shovels to rapidly relocate flexibility the capability to work in faces of different heights or to dig at different levels of a face is an advantage wheel loaders are most productive at face heights of at least three times the bucket height although at shorter face heights they can drive forward through unconsolidated or wellbroken material with only slight impact on productivity in lessconsolidated material hydraulic shovels can penetrate a bank at different levels to separate material at the face with only slight impact on productivity most hydraulic shovels can also be configured as backhoes or mass excavators for digging below grade and loading trucks below or at operating level loading trucks below operating level allows use of a number of spotting techniques and shortens cycle times loading trucks at operating level although slower is obviously desirable when the pit bottom is wet backhoes are generally limited to use where face heights are about equal to the stick length mining shovels require a face height of about 50 of their point sheave height basically about the height of the teeth when the dipper stick is horizontal however they can operate at higher bench heights than can wheel loaders and hydraulic shovels which reduces other operating costs a higher face means fewer benches fewer relocations and lower drill andblast costs life economic lifetimes for loading tools are generally as follows wheel loader 5 to 7 years hydraulic shovel 7 to 10 years mining shovel 15 years of course with enough replacement components the life of any unit can be extended longer life is arguably a desirable feature it is certainly necessary to justify a higher capital cost but a system with a shorter life but lower operating costs and high resale value can be an equally good or better choice support requirements several factors affect support requirements for a loading tool drive system large mining shovels are currently available only with electric drives wheel loaders and hydraulic shovels generally have diesel drives although very few wheel loaders and some hydraulic shovels are optionally available with electric drives electric drives have lower and more consistent operating costs but require in pit electrical reticulation systems comprised of electrical substations and power distribution cables involving specialized support equipment personnel planning and operations diesel drives require fuel transport for refueling in the field but the equipment involved is usually common to mining operations using large equipment digging profile wheel loaders and hydraulic shovels more so than mining shovels can flat pass and thus require minimal cleanup assistance however mining shovels with their greater reach can stand back farther from the face and so keep the truck back farther from the toe a

lthough one could argue that this class of loading tool should not waste time doing cleanup work that is better left to auxiliary equipment maintenance larger hydraulic excavators and mining shovels have limited mobility and transportability so all maintenance on these machines must be conducted in the field material conditions the condition of the floor and bank affects the different types of loading tool differently a wet or soft floor causes traction problems for wheel loaders which can significantly impact tire costs and productivity for soft floors track shoes can lower the ground bearing pressure exerted by a shovel for very soft floors extrawide track shoes can minimize the ground bearing pressure although they reduce maneuverability due to the increased turning forces required operating methods the most common loading tool operating method is one sided loading figure 10 3 17a also called single sided loading this method requires minimal pit support coordination and minimizes variations in traffic patterns and truck movements the latter of which often significantly influences safety the method also has a relatively small footprint so it can easily be implemented in benches only 30 40 m 100 125 ft wide in the most common variation of the method the truck stops or queues in a position where it can observe the loading operation which allows the truck operator to ensure that the area is clear after the previous truck has been loaded and moved off the location of the queuing and reversing points is generally left up to the truck operator as it changes so frequently however choice of location can significantly influence cycle time so training on efficient techniques is important road constrictions pit obstacles poor spot selection cleanup and cable handling are common causes of unnecessary delay traffic patterns are best designed to allow the truck to reverse with the shovel on the truck operator s left and thus always visible to the operator either directly or in the left side mirror one sided loading has several disadvantages mining shovels and hydraulic shovels cycle in 30 35 seconds and spotting takes 45 60 seconds so the shovel must wait for the truck reducing production levels in addition the time required for a cleanup dozer to work in the spot area can delay operation these disadvantages can be addressed to some extent by two sided loading figure 10 3 17b also called doublesided loading which reduces the shovel delay between trucks although this method appears to be just a doubled up singlesided loading method it actually adds some complexity the road circuit is more complex because it requires a y intersection a much larger working face up to three queuing points and twice the cleanup and road maintenance support electricpowered loaders also require a cable bridge and have more confined turnaround areas all of these factors dictate a larger footprint for the loading area so the

method is more suitable for benches at least 80 m 250 ft wide furthermore near the ends of the working face the work area becomes restricted and the shovel generally must revert to single sided loading so two sided loading favors wide working faces two sided loading can increase production by 5 10 with the general caveat that the more passes per truck the less the benefit since improvements come from reducing time between trucks wheel loaders spot trucks somewhat differently than do the other types of loading tool figure 10 3 18 from a wheel loader trucks are best spotted at a 45 angle which allows the loader to approach the face and the truck at right angles with minimal turning between the two a good setup requires less than one tire rotation of travel distance in each direction to load the truck a less common method drive by loading figure 10 3 19 is used with bottom dump trailers or backhoe loaders it does not require reversing rather the truck merely drives alongside the loader it requires a very narrow face with a long clear bench adjacent so that the truck can approach along the bench and the loader need never move far from the bench backhoes offer a few variations for loading trucks on the floor because the truck can back to the face at virtually any angle the truck spot can be set up to bring the material over the side or through the tailgate depending on the bench configuration some hydraulic shovels without independent pumps for hoist and swing can benefit from positions that emphasize either hoist or swing rather than working both together matching loading tools to haulers selecting the best loading tool size to hauler size involves analyzing the number of passes required to load a truck and the number of trucks needed to match the shovel the primary goal is to optimize the total loading and haulage cost maximum production is determined by the loading tool not the hauler however pit production costs are most significantly influenced by the hauler figure 10 3 20 specifically haulers account for nearly 50 of the total system cost and loading tools only about 10 thus the following bears repeating in general the loading tool drives production and the hauler drives cost the strategy in matching loading tool size to hauler size is to consider but not be ruled by the concept of matching passes and minimizing the number of passes per truck the variety of loaders and trucks available make it virtually impossible to always achieve a perfect match another complication is that variations in material density and bucket fill ensure that no two dipper loads will be exactly the same in fact the distribution of dipper and truck load sizes is nearly normal this distribution pattern is considered in loader design and some latitude exists in matching dipper loads to truck loads target payload should be changed only after consultation with the manufacturer given that spot time cannot be less than

shovel cycle time it follows that the more passes per truck the more the shovel and therefore the system produces the downside of this premise is that it favors selection of smaller trucks which of course means higher operating costs and more congestion the premise is therefore best considered as suggesting that truck size not shovel size determines the number of loading passes in reality variations in production or fleet costs caused by pass match are not as significant as the consequences of undertrucking or overtrucking a fleet undertrucking a typical fleet by only one truck offsets the production advantage of the additional pass per truck likewise overtrucking a fleet by only one truck offsets the cost advantage of a larger truck with fewer passes per truck track dozers large track dozers figures 10 3 21 and 10 3 22 are extremely common in all mining operations they are designed to move the greatest amount of material in the most efficient way and generally used for both utility and production work utility work includes tasks that support a mine s main production fleet such as dump site preparation and cleanup bench preparation road creation stockpile work and reclamation however the focus of this section is their use for production work specifically mass excavation for which they excavate push and rip in situ or blasted material from one area to another examples of this use include assisting primary production loaders and slot dozing figure 10 3 23 for purposes of comparison the track dozer is a lowcapital cost high operating cost system that has moderate flexibility and can operate through a limited range of applications with moderate sensitivity to geologic variance currently two major manufacturers build track dozers for the mining industry caterpillar and komatsu table 10 3 3 compares their specifications for typical large track dozers used for mining production the specifications are very similar between manufacturers except for the komatsu d575 super dozer sd whose high end size class is an anomaly most of the industry uses the smaller size classes because of their lower operation costs and flexibility track dozers are complex machines because of their variety of mechanical electrical and hydraulic systems all fitted into a compact design that protects against the elements operating conditions are unlike those for other materialexcavation equipment that loads statically and advances as the face moves rather slot dozing and rip and push operations to assist large loaders require the dozer to push material at varying distances and gradients in poor ground conditions therefore machine design must be very robust to this end the mainframe is rigid and consists of multiple fabrications and castings all major components and systems are mounted to the frame the radiator engine torque divider transmission brakes and final drives are housed in the frame and body and in most case

s are modular for ease of removal and installation the components that typically require extensive forethought for maintenance are those that engage the ground commonly referred to as ground engaging tools gets gets include blades and wear plates rippers if fitted rollers and idlers and tracks all dozer manufacturers provide a large variety of options for these components except for rollers and idlers the blade and associated wear plates are customized for each application for instance caterpillar provides five types of blade semiuniversal universal reclamation coal and carry dozer two types of ripper single shank and multiple shank are available and both regular size and wide track shoes are available the goal is to maximize get life and productivity by matching the machine configuration and options to the site characteristics thorough site analysis by the manufacturer is required to evaluate the production cycle and material characteristics the longer the push distances potentially the higher the replacement frequency material characteristics such as fragmentation size abrasiveness and cohesiveness must be evaluated to determine a dozer s replacement life dozer productivity the first step in determining track dozer productivity is to calculate how much material a particular dozer can push this narrows the number of suitable size classes and defines some viable configurations before detailed productivity calculations are made the following factors are involved in determining dozer capability weight the dozer cannot push more than its weight coefficient of traction this is the percentage of the dozer s weight that can be pushed for a given material before the track shears or slips for most materials this value averages 60 for loose sandy material it can be as low as 30 multiplying the coefficient of traction by the dozer s weight gives the weight that a dozer can push material density the denser the material the smaller the volume of material that the dozer can push blade capacity and selection are directly related to material density carry force ratio this is the energy needed to compensate for friction or drag and push material across itself it is usually about 10 and is included in calculations of drawbar pull which provides the corresponding machine speed slope this is the percent grade downhill or uphill it is included in calculations of drawbar pull which provides the corresponding machine speed push distance this is the same as the return distance although push and return are traveled at different speeds both have a large impact on cycle time in production dozing it is critical to plan dozer size and configuration so as to maximize productivity a common tagline in the industry is big load slow which describes how to manage a machine for maximum productivity productivity is defined as work done over a unit of time the industry continues to debate whether do

zer productivity is higher when pushing larger loads slower or smaller loads faster table 10 3 4 shows a case study of two identical dozer configurations for a slot dozing application cycle conditions are identical and the average push distance and gradient are approximately the same the average cycle time is about 10 slower for dozer 1 than for dozer 2 in large part because the blade load is 25 larger the end result is that productivity is 5 higher for dozer 1 pushing larger loads slower among the factors that can improve the operating efficiency of a dozer are technique and technology technique has to do with how an operator approaches a particular job the following technique tips reflect best practices for a variety of dozer operations related to production dozing and ripping where the highest productivity gains can be made in some cases these adjustments can increase productivity by up to 25 slot front to back technique most efficient operator works the cut from front to back push distance increases with each pass efficiency is optimal due to downhill blade loading creates the slot and uses it throughout the cut slot back to front technique less efficient operator works the cut from back to front push distance decreases with each pass efficiency suffers from uphill blade loading does not fully use the slot throughout the cut slot back each pass technique less efficient operator starts each pass at the back of the cut each pass uses the entire length of the cut at a uniform depth efficiency and productivity suffer because the machine travels the entire length of the cut in both directions with each pass berm criss cross removal most efficient for removing center berms operator works the cut from back to front push distance decreases with each pass existing slots are used to hold in material and increase blade load berm management berm should not exceed blade height and should be high enough only to trap material for optimum loading center berm width should be one third the blade width for optimum productivity the smaller the berm the easier the disposal ripping operator should rip downhill when possible operator should reduce speed in shock and impact conditions operator should try cross ripping if material does not free up operator should pull the ripper tip forward after penetration the final piece to improving dozer efficiency concerns technology dozer manufacturers have developed software features that increase safety and efficiency these features provide automatic control or even real time data to allow the operator to adjust to conditions one such feature is caterpillar s autocarry caterpillar 2008 which automatically optimizes the blade lift and lower functions during the carry segment of push dozing by monitoring calculating and integrating data on power train output ground speed track slip and tractor attitude another

such feature is an onboard global positioning system gps that links directly to survey data in the mine office the operator obtains needed data via a cab mounted display rather than by reading surveying maps or looking for grade stakes gpss are available from the major manufacturers and from third party suppliers haulers the hauling of material such as coal ore sand gravel or topsoil from one point to another in a safe efficient and costeffective manner is critical for mining operations selecting the type of surface hauler requires a thorough understanding of the selected mining method and its associated advantages and disadvantages along with the machinery available to the industry a number of haulage options are available for mining all of them with unique characteristics that can be optimal for a particular mine site and haul distance figure 10 3 24 compares haul distances for dozers front end loaders wheeltractor scrapers articulated dump trucks off highway rigidframe trucks and belly dump haulers each of which has an economic advantage at certain haul distances with some overlap on highway trucks on highway trucks are not widely used in mining operations because of their lack of capacity which is only 6 12 m3 8 15 yd3 and their limited performance capability however some mine sites use these trucks for hauling coal eastern united states prestripping ttopsoil in contract surface mining operations australia and hauling construction aggregates for road building their capability to travel long distances at relatively low cost while meeting local on highway regulations and their overall flexibility in mining operations where landscape is limited provide a unique hauling alternative three types of on highway truck are commonly used for mining operations dumptrucksguide com 2006 1 standard dump truck this truck typically has a two or three 1 front 1 2 drive axle truck chassis with a dump body mounted on the frame the dump body is hoisted hydraulically by cylinders mounted between the cab and the front wall of the body the small size of the truck allows for exceptional maneuverability in tight loading areas 2 semitrailer rear dump truck this tractor trailer combination typically has a three axle tractor and a two axle trailer the trailer body is hoisted hydraulically key advantages compared to a standard dump truck are faster unloading and increased payload 3 semitrailer belly dump truck this tractor trailer combination typically has a three axle tractor and a twoaxle trailer with a c shaped dump gate the dump gate mounted on the trailer is hoisted hydraulically a key advantage is the capability to unload material as a wind row the advantages of on highway trucks in typical mining operations are few but there are applications for which they can provide an adequate hauling alternative their small size allows for flexibility in operations where loading areas are small as for eastern u

s contour coal mining operations their low fleet cost is advantageous when mining operations need additional hauling capacity for short term use they require substantially less investment in up front capital and subsequent operating cost than do large off highway mining trucks they can also be used as a secondary fleet for special projects such as road construction and prestripping which are typically contracted from local construction firms the disadvantages of on highway trucks are based primarily on their relative performance compared to that of large off highway trucks payload capacity is an obvious difference on highway trucks are smaller and thus have higher overall fleet costs per ton of material due to their lower productivity and the larger number of trucks required additionally they are not designed for use in rigorous mining conditions their structural design electrical and hydraulic systems brake and steering performance and power train are designed for highway use not for 24 7 mining applications in mining operations haul road gradients plus rolling resistance can be as high as 20 on highway trucks can have difficulty in deteriorating road conditions to the point where support equipment is needed to recover them reducing fleet productivity thus they are not considered to be a primary haulage solution for the mining industry although they will always have limited use in mining applications off highway articulated trucks off highway articulated trucks are a hauling alternative primarily for middle to large scale construction projects they are often used for prestripping road construction and material hauling for ground preparation for buildings and other infrastructure when used with small hydraulic excavators or wheel loaders they can constitute an effective loader and hauler fleet for a mass excavation project they are widely used for pre and postmining construction in soft underfoot conditions in small loading and dump areas or on steep 10 15 grades an articulated dump truck adt figure 10 3 25a is a three axle machine with an articulation point between the front axle and the two rear axles the articulation which is unique to this truck type is useful where there is limited area in which to operate the three axles all provide power to ground this all wheel drive capability provides an advantage over on highway or off highway rigid trucks whose one or more rear axles are the only source of power to ground it also allows the adt to operate well in soft underfoot conditions defined as rolling resistances of 10 20 caterpillar 2000 an adt can vary power to the wheels according to road and haul conditions for example a caterpillar adt has three operating modes a standard mode 40 60 split between front and two rear axles a low mode 50 50 split and a high four wheel drive mode all three axles have equivalent power to ground the dump body is mounted on the rear frame with tr

aditionally two options to dump material the first option is similar to that for on highway dump trucks two hydraulic cylinders hoist the body dumping material rearward the second option is an ejector body figure 10 3 25b with a hydraulically moveable front wall that runs on a rail system fixed to the side wall the front wall pushes material back and dumps it rearward despite its added design complexity an ejector body increases productivity by decreasing both dump times and the amount of carryback per load it also allows the machine to safely dump in steep inclines or side slopes a number of major global construction and mining manufacturers provide adts for the mining industry including caterpillar john deere komatsu terex and volvo in addition a number of regional manufacturers in china india and elsewhere have offerings adts generally range from 23 to 36 t 25 to 40 st in four size classes that increment every 4 5 t 5 st table 10 3 5 highlights their specifications by size class payload adts providing the surface mining industry a hauling solution that is adaptable to tough hauling conditions however because of their limited hauling capacity they are not viable as primary production machines off highway rigid frame trucks the primary hauling machine in mining is the large offhighway rigid frame truck figure 10 3 26 in the 1950s and 1960s it was used as an alternative to locomotives and small dump trucks over the years it has proven to be cost effective flexible across a variety of applications and capable of handling the rigors of 24 7 operation it continues to be a frequent hauling solution of choice for surface mining an off highway rigid frame truck has a rigid unarticulated frame constructed of multiple steel fabrications and castings to this frame are mounted the truck cab body diesel engine power train front and rear suspension front and rear wheels tires and more all interconnected by sophisticated mechanical and electrical hardware and software the two most distinguishing characteristics of the various makes and models are payload capacity and power train type payload capacity the size or payload of an off highway rigid frame mining truck plays a significant role in determining the viability of a mining operation for example for a given mine plan one could elect to operate either fifty 90 t 100 st trucks or twenty five 180 t 200 st trucks at first glance assuming that machine performance is equal the second choice at half the fleet size would seem to reduce costs dramatically however each fleet carries a different capitaland operating cost footprint and these footprints should not be assumed to be linear trucks for surface mining currently have payload capacities of 90 360 t 100 400 st these values have evolved over time driven by the mining industry s desire to go larger in order to maintain or increase production while decreasing fleet size and operating co

sts within this payload range there are five distinct classes designated according to size 1 90 t 100 st class 2 135 t 150 st class 3 180 t 200 st class 4 220 t 250 st class 5 290 t 320 st ultra class truck uct figure 10 3 27 shows current mining truck models by size class and manufacturer some manufacturers are not included because information was not available product strategies regarding payload capacity clearly vary by manufacturer most mining truck manufacturers work closely with the industry to determine the appropriate size for an application typically matching the rated payload capacity to the current and expected future loading tools using three to five passes as the optimal level for example for a truck with a rated payload capacity of 220 t 250 st an electric cable shovel with a capacity of 46 m3 60 yd3 assuming a 90 fill factor could load 1 780 kg lcm 3 000 lb lcy of material in approximately three passes as important as payload is to a mining operation the following points continue to be debated can we get more payload are our truck payloads at optimal levels should we upsize some of these questions can be answered by considering the loaders and associated practices but in some cases the trucks should be considered as well thus after a decision is made about size class the next decision should concern the type of truck body there are now a myriad of truck body designs for any type of mining truck and the choice of design depends on the material characteristics of the mine which differ from country to country truck bodies today can be customized for each operation so as to maximize payload reliability and durability for the purposes of such customization a mine operation creates a profile listing the following information material type material density lightest needed to ensure that payload is met regardless of any fluctuation in material density material fragmentation size needed to determine linerplate thickness and dumping characteristics material abrasiveness needed to predict wear characteristics for determining the appropriate liner package material cohesiveness also needed to determine the appropriate liner package loading tools model and type percent utilization bucket size body mining truck dimension limitations needed to determine maintenance facilities hoppers crushers loading tool dump heights and more another point of discussion is payload management the constant push to increase payload for an existing truck fleet is not a bad thing if trucks consistently perform on average at under their rated levels because pushing trucks beyond their rated levels can be detrimental most manufacturers have a payload policy that outlines the levels at which a truck can perform within the certification or design envelope of the machine standard practice today is a so called 10 10 20 payload policy figure 10 3 28 that distributes truck

payloads over a set period of time to address risks associated with overloading the highest risk is overloading beyond 120 of target payload all mining trucks are designed to meet certain society of automotive engineers sae and international organization for standardization iso design standards that address not only component and system functionality but also safety two particular standards are related to payload iso 3450 1996 concerning brake certification and iso 5010 1992 concerning steering certification the requirement for certification is that a truck loaded to its maximum payload an overloaded state at the upper end of the normal distribution of payloads should stay within requirements overloading beyond the payload policy risks among other things machine durability as the structural components are designed to certain life targets and if overloaded more frequently than recommended become fatigued and prone to fail early failure increases maintenance and repair costs dramatically due to unplanned rebuilds and repairs payload management can be difficult to apply in practice but it plays a large role in meeting production requirements and keeping a truck fleet operating safely and efficiently it is important to check with the manufacturer for the tested limit specific to a truck finally when should a truck fleet be upsized to a new size class moving to a higher size class has many benefits and at least a few barriers which become more daunting with increasing truck size the following is a list of considerations caterpillar 2009 mine operations mine design larger mining trucks can require changes in haul road design and load and dump area especially when upsizing to ucts loading tools it is important to have the proper loaders to meet production requirements with the new fleet support equipment larger mining trucks place a stronger demand for support equipment to maintain haul roads and loading dump areas operator training curriculum and training tools must be changed mine maintenance facilities maintenance facilities such as shop lube islands and parts and component storage may need to be upgraded to handle the larger machines tooling tooling requirements must be upgraded to handle larger components and any specialized tools training significant training may be required when upsizing regardless of system commonality drive train type truck performance is the next key piece of the puzzle three types of drive train are currently in use 1 mechanical drive similar to those used in on highway automobiles and trucks 2 direct current dc electric drive 3 alternating current ac electric drive figure 10 3 29 shows a side by side comparison of the mechanical and electric drive trains the mechanical drive train contains five major components engine torque converter transmission differential and planetary gear sets wheels the power source is the diesel engine plus t

orque converter the latter transmits rotational power from the engine to the main driveshaft the transmission controls machine torque and speed during operation the differential transfers output torque to the wheels two hydraulic brake packs are mounted on each of the axle shafts on the two rear wheels the entire system is activated by means of a variety of electronic control modules and hydraulic control systems dc and ac electric drive trains contain six major components engine generator power converter wheel motors planetary gear sets wheels and retarding grid the power source is the diesel engine plus generator the latter converts mechanical power from the engine into electric power ac current from the generator is then converted into useable form in a dc drive truck a rectifier converts it into dc power in an ac drive truck a rectifier converts it into dc power and inverters convert it back to a controllable version of ac power suitable for managing the amperes volts and frequencies of the wheel motors in order to create machine speed and torque the dc or ac wheel motors receive the electric power and feed it mechanically to the planetary gear sets wheels the retarding grid a bank of resistor elements provides braking force by turning the wheel motors into generators creating power rather than receiving it this power is sent through a control cabinet and on to the resistor elements the resistors impede the flow of the electric power which causes the wheel motors to slow rotation heat generated is cooled by an electric fan the three types of drive train have unique performance characteristics that impact productivity and operating costs the major points to compare are the following examined in more detail in table 10 3 6 system limitation on grade speed on grade propelling speed on grade retarding top speed fuel consumption maintenance and repair costs other systems these additional systems deserve serious consideration bottom dump coal hauler trolley assist mining truck bottom dump coal hauler a semipopular haulage solution for surface coal is the bottomdump coal hauler it has become a staple in thermal coal operations because of the need to haul coal from the pit to a nearby power plant the truck trailer configuration figure 10 3 30a of the bottom dump coal hauler has a mining truck chassis as the tractor modified with a hitch assembly to receive a trailer the most common size class is the 90 136 t 100 150 st standard truck chassis the high volume trailer has typically 1 5 to 1 7 times the payload capacity of the corresponding truck fitted with a rear dump body this high capacity is well suited to long hauls with few high gradient segments and the bottom dump coal hauler provides a higher production rate than does a traditional mining truck thus lowering haulage costs per ton for the truck fleet the main players in this market segment for the chassis are caterpillar 777

and 785 models hitachi ch120 ch135 and ch150 models and komatsu 785 and hd1500 models all of whom provide the necessary chassis modifications from the factory the trailers are designed and manufactured by smaller specialty firms such as kador engineering australia maxter atlas canada and mega united states these firms provide the trailer and hitch assembly whereas truck oems provide the additional axle to be fitted to the trailer in the interest of product consistency one firm provides the complete package rimpull united states provides an entire lineup of tractor trailer options with their cw160 cw180 and cw200 models the truck trailer configuration of the bottom dump coal hauler has limited application due to the design specifications of the chassis a fully loaded rear dump truck has a continuous rating that allows for total effective gradients of 10 15 when the additional weight of a trailer is added and payload is increased performance drops limiting the machine to effective gradients of just 5 10 this is suitable for a number of surface coal operations especially where the coal seam is relatively shallow and there is a limited amount of cover or overburden another point of limitation is haul distance the truck trailer configuration is economically viable only for haul distances of 3 2 km 2 mi for shorter distances this configuration with its added costs associated with the trailer tires higher fuel consumption due to an increase in cycles and lower performance characteristics does not compete well against a traditional truck configuration however for one way haul distances of 8 16 km 5 10 mi all of the benefits associated with this configuration outweigh those for the rear dump truck a variant of the truck trailer version of the bottom dump coal hauler is the unibody configuration or unibody coal hauler figure 10 3 30b currently manufactured only by kress united states this machine has a built in bottom dump body i e it is unitized among its advantages compared to the truck trailer configuration it has a significantly higher payload to weight ratio a higher horsepower to weight ratio and a 50 t 55 st lower empty weight it also has higher fuel efficiency fuel consumed per ton and lower metric ton kilometers per hour ton miles per hour which improve tire life its higher horsepower to weight ratio can enable it to for example operate at higher effective gradients to increase productivity in addition its drive train is capable of achieving higher top speeds typically up to 30 higher than is the drive train for the tractor trailer configuration which is limited to the speed capability of the chassis this can increase productivity measurably on long hauls trolley assist mining truck a trolley assist mining truck figure 10 3 31 is a unique application for mining trucks and strictly exclusive to electricdrive models its use for large scale material transp

ort dates back to the late 1930s in italy s full trolley systems its use for mining began during the energy crisis of the 1980s with upgrades in technology it still has relevance in the mining industry currently two suppliers offer a trolley assist on their mining trucks hitachi and komatsu four operations use it today all in africa a trolley assist mining truck draws its power from overhead power lines that are run on haul segments where the largest benefits can accrue such as where the loaded truck operates on a positive grade the truck is fitted with a pantograph that acts as a conduit between the line and the truck s electric drive distribution system as the truck approaches the line the operator lifts the pantograph until it contacts the line when the two engage the operator removes his or her foot from the throttle and continues to steer while the truck draws power from the line power is fed to the wheel motors temporarily replacing the diesel engine and generator trolley assist has several benefits decreased fuel consumption achieved by running the engine at idle for the length of the line depending on the length of the ramp fuel savings can be as high as 50 increased productivity per cycle achieved by using the excess power capacity in the wheel motors the power rating of the wheel motor is almost twice the engine gross horsepower in order to meet the technical requirements for continuous operation under diesel power that is the primary reason for the significant speed on grade performance in retarding when the wheel motors use their full potential the same principle applies during trolley assist when power from the diesel engine is replaced by power from the overhead line the result is an increase of up to 80 in speed on grade performance depending on the haul cycle this can translate into an increase in production of up to 10 increased diesel engine lifetime the heaviest toll on an engine in a haul cycle normally occurs when the truck is fully loaded on a grade the very point at which trolley assist kicks in the life of a diesel engine is calculated in terms of the total quantity of fuel consumed during the design life with the engine operating at idle on grade rather than at maximum the life can be extended significantly potentially eliminating one complete engine rebuild over the life of the truck trolley assist also involves additional operational costs and constraints including the following relative costs of electric power and diesel fuel this is one of the single largest variables to consider when evaluating trolley assist the cost for diesel fuel can be enormous for a medium to large fleet but the ultimate question is whether the savings in fuel can offset the cost of electric power capital cost of trolley wayside equipment this consists of mine power distribution substations masts and wire capital cost of truck trolley equipment this consists of a pantograph aux

iliary cooling and truck controls mine plan trolley assist does not allow for operational flexibility after the equipment is in place it typically is not moved until doing so makes economic sense often 5 to 10 years from the initial installment therefore it is critical to evaluate the long term mine plan and determine whether or not a permanent main haul road is possible haul profiles determining which haul cycle benefits from trolley assist is one of the most critical pieces to the evaluation a long haul segment with a grade that the truck travels loaded is the best choice capital cost for additional mine support equipment costs associated with additional motor graders or wheel dozers may need to be included haul roads where trolley assist is used must be kept in pristine condition since spillage and rutting can cause the truck to lose connection with the overhead line the industry will continue to support trolley assist technology improvements now under consideration include concepts such as auto control when ascending a grade and regeneration of power when retarding during a return cycle wheel tractor scraper one of the oldest concepts of bulk material handling is the wheel tractor scraper wts figure 10 3 32 which traces its roots back to horse drawn slip scrapers in the late 1800s asme 1991 today the wts is the only machine that can self load haul and dump with a single operator for purposes of comparison the wts is a low capital cost high operating cost system that is very flexible and can operate through a limited range of applications with high sensitivity to geologic variance mobility and flexibility are key characteristics of the wts which makes it ideal for small short life mining projects its capability to remove and place material in controlled lifts makes it the machine of choice for topsoil relocation in reclamation operations wtss are available with three types of loading pan elevator and auger the pan uses the motion of the machine to force material in to the bowl the elevator and auger have mechanical apparatus that assist the material into the bowl the pan is slower to load but is better suited for blocky materials the largest units have a bowl capacity of 34 m3 44 yd3 for earth and rock densities and can be larger for lighter materials such as coal wtss are either single engine or twin engine systems that can be pushed usually by a dozer to assist with loading larger units can also configured in a push pull arrangement for connecting two scrapers during loading thus putting the horsepower of two machines on one cutting edge because of its cutting mechanics the wts best suited to unconsolidated materials production is obviously affected by haulage distance but for a moderate haul of 450 m 1 500 ft a large wts can produce at a rate of about 300 bcm h 400 bcy h it is capable of speeds 50 km h 30 mph optimal one way haul lengths are 200 1 200 m 400 4 000 ft t

he basic haul cycle is similar to that for a truck load haul dump and return because the load and dump components of the cycle can be 50 100 m 150 300 ft long it is most efficient to set up the haulage route so that the loaded haul is shorter than the return the effect of haul distance on production rate is not quite 1 1 for example doubling a mid range haul distance decreases the production rate by about 40 properly designed and wellmaintained roads are as critical to a wts as to any other hauler although this fact is often overlooked possibly because the load and dump areas are relatively rough a smooth haul not only lowers rolling resistance it also makes for a smooth ride with reduced loping and therefore higher speeds under most conditions wts load times are in the range of 30 60 seconds production can be enhanced by downhill or assisted loading figure 10 3 33 a dozer assist or a push pull system delivers additional horsepower increasing the production rate by about 10 having a ripper equipped dozer handy can help to loosen hard packed material between scraper loads loading is accomplished by lowering the bowl until the material flows steadily cutting too deep may take longer resulting in higher fuel consumption and spinning tires with their associated costs wts dump or spread time is about 20 30 seconds and the total fill time is 35 45 seconds caterpillar 1998 a significant feature of the wts is its capability to spread a load in a controlled manner laying material down in thin lifts improves compaction on multiple lifts or allows material to be spread to a specified thickness in pit crushing and conveying system a relatively recent and increasingly popular entry into large surface mining systems is the in pit crushing and conveying ipcc system figure 10 3 34 this system uses a crusher sizer unit to process material from a cyclic loader to a size that is suitable for conveyor transport extending the application of around the pit conveyor systems to include consolidated waste and overburden for purposes of comparison the ipcc is a high capitalcost low operating cost system that has limited flexibility and can operate through a limited range of applications with moderate sensitivity to geologic variance in pit or near pit crushing of ore has always been common primarily because the location of the crusher station has only a very small impact on the total comminution costs crushing of waste on the other hand until the advent of this technology has been difficult to justify on the basis merely of enabling greater use of conveyor transport although ipccs have been used since the 1950s the first large scale mining systems capable of working with feed direct from a mining shovel were implemented only in the 1980s with the implementation of compact twin roller sizers in about 2002 the system finally gained wider acceptance the ipcc is a mobile crusher sizer machine historically mobile mean

t relocatable usually with significant cost infrastructure and time requirements however the ipcc is self propelled or easily transportable with no fixed infrastructure requirements and travels with its own loading tool for crushing the ipcc crushes material to a size suitable for feeding to a conveyor belt in all three dimensions the material can be no more than about 30 of the width of the belt and a significant percentage of the material must be smaller than that in order to cushion the belt and prevent damage to it although the ipcc s crushers and sizers can handle materials with extreme rock strengths to date the ipcc for waste is most cost effective for use with materials whose compressive strengths are 50 90 mpa 7 000 13 000 psi and somewhat less cost effective for harder or softer materials for conveying the ipcc relies on a series of components that feed material from one to the next a cyclic loading tool such as mining shovel feeds into a mobile sizer that follows along with the shovel the sizer feeds into a face conveyor much as for a bwe the face conveyor feeds into a series of other conveyors eventually leading to a discharge conveyor at the dump because of the multiple conveyors the ipcc is suitable for pit geometries that favor use of draglines and bwes i e those with long linear faces because its loading tool is a shovel it can handle materials with rock strengths greater than a bwe can handle at the time of this writing however ipcc systems are still finding their niche application but show favor for waste transportation in linear pits with geometries that require long hauls with relatively limited elevation change the digging face of the ipcc has a large footprint similar to the case for the bwe which needs to be considered when planning other pit operations such as blasting or pit access conversely the conveyor route has a relatively small footprint which can reduce costs for ramps earthen bridges and the like likewise the dump area can be configured with some flexibility and material can be placed in final form with very little rework required for rehabilitation the ipcc has an electric drive system although this type of system requires electrical infrastructure in the pit it does not depend on diesel fuel whose costs tend to fluctuate independent of the product market it also does not have tires certainly the tire shortage of 2006 2008 had a significant influence on a number of ipcc purchases the ipcc generally requires minimal operating labor per unit of production operational and maintenance labor requirements are cyclic with significantly higher demand during conveyor relocation these cycle demands are relatively easy to manage where maintenance contractors are available the operating method lends itself to semiautonomous operation of some components further decreasing labor demand conveyor relocation these cycle demands are relatively easy to manage where mainte

nance contractors are available the operating method lends itself to semiautonomous operation of some components further decreasing labor demand tubs to provide lower ground bearing pressures and various walking mechanism systems have been developed over the past 30 years to provide a longer life at a lower operating cost a dragline with a 170 m3 220 yd3 bucket a 122 m 400 ft boom and a mass of 12 700 t 14 000 tons had been in use in the united states however because of environmental considerations most of the largest draglines particularly in the united states have been removed from service good bench preparation for walking draglines requires that the material be well blasted for good bucket loading yet remain sufficiently stable for the dragline to operate and move without the highwall collapsing it is common for the dragline bench to be prepared so that any point loading that can damage the tub is kept to a minimum the costing of equipment such as tracked bulldozers used for this bench preparation needs to be included when calculating system costs stripping shovels a stripping shovel with a bucket size similar to a dragline will be more productive because of its crowd and breakout ability shorter cycle times and its ability to handle dense rocks although a competent floor is necessary and may require some preparation stripping shovels do not require the more extensive bench preparation necessary for draglines so this additional cost is saved the largest stripping shovel used to date had a 138 m3 180 yd3 bucket with a 65 m 215 ft boom and installed power of 22 500 kw 30 000 hp as with draglines the largest stripping shovels have been taken out of service in favor of truck shovel operations largely because of environmental considerations size selection the primary steps in selecting a cyclic excavation machine involve 1 bucket capacity selection 2 determination of machine geometry and 3 reassessment of the first two steps into a standard model the first step in loading tool selection should be to determine from the mining method planned the planned production rate in whatever units are applicable and then convert these units into a loose volume to be moved per hour table 10 4 1 shows applicable factors for material densities and swell factors for a list of common materials this will allow the calculation of the required bucket size using the following equation 3 600 q b fe a p t 10 4 1 where q is the bucket dipper capacity p is the required production loose volume per hour t is the theoretical cycle time bf is the bucket fill factor e is the job operating efficiency and a is the mechanical availability expected over the period of operation theoretical cycle time values for theoretical cycle time t can be obtained from information supplied by the oem or from time studies of similar machines in similar conditions table 10 4 2 provides average cycle time for a range of

bucket capacities for draglines and stripping shovels bucket fill factor the bucket fill factor bf is a factor of the material sizing condition and the ease or difficulty of filling the bucket ideally this factor is best determined by field measurements however if this information is not available typical bucket fill factors can be found in table 10 4 1 job operating efficiency job operating efficiency e is the percentage of time that a machine will actually operate as dictated by maintenance scheduling and operating practices good management of the operation typified by excellent management and supervision planned maintenance programs high availability and so forth will result in a better operating efficiency conversely poor management as typified by sloppy maintenance practices low machine availabilities and so forth will result in a lower operating efficiency average operating efficiency figures should range between 0 75 and 0 90 0 83 average mechanical availability mechanical availability a is the percentage of time that a machine is running and available to work in good operating conditions with good maintenance practices availability will be higher as conditions deteriorate extreme temperatures dusty conditions etc and maintenance practices become poorer availability will be reduced availabilities of 85 to 95 should be achievable in most operating conditions during at least the first few years of equipment operation one further adjustment that needs to be made for cyclic excavation machines is to divide the calculated bucket capacity by a propel factor to make allowance for time required for the machine to move the typical propel factor for a walking dragline is 0 94 a typical propel factor for a stripping shovel is 0 96 machine geometry now that the bucket capacity has been selected the next step is to consider machine geometry primarily the dumping radius and the dumping height normally the cut width is equal to the pit width and all exposed material is loaded out in a typical dragline pit the minimum cut width is determined by the mineral loading and transport equipment requirements for small stripping shovels cut width can be as narrow as 15 to 18 m 50 to 60 ft with larger shovels requiring a pit width of 24 to 30 m 80 to 100 ft in practical considerations a narrow pit width allows more efficient use of the spoil space and reduces the dragline cycle time allowances will also need to be made if the overburden to be removed has widely variable thickness the geometry of cyclic excavation machines can be detailed more effectively by drawing or computing average and extreme pit sections and plans showing both the stripping and loading operations multiple iterations can be done to finalize the geometry final selection after bucket size and operating geometry dumping radius and dumping height are determined the optimal machine can be selected from oem literature although i

t may be impossible to match exactly with calculated numbers the nearest model can be adapted for example a reduction in dumping radius will result in a reduction in dumping height and vice versa because these units often involve customization the mining company should work closely with the oem on final design specifications cost estimation figure 10 4 1 can be used to calculate the estimated hourly ownership and operating costs for the various machinery and equipment referenced in this chapter based on information available from the oems and or actual experiences in similar applications bucket chain excavators a bucket chain excavator bce excavates material below the grade of the main house unit and transports it upward and away from the unit it can mine high outputs in weak unconsolidated ground its primary advantages are its excellent downward digging ability and its moderate ability to dig upward its greatest disadvantages are its inability to dig hard ground or to excavate materials selectively bucket wheel excavators bucket wheel excavators bwes are very effective for mining large volumes of unconsolidated material although they also offer some productive ability in harder formations they are also more selective than a bce and can accurately cut bands as narrow as 100 mm 4 in albeit at a lower output one primary disadvantage is their inability to achieve downward digging action without special modifications the bwe digs with a series of evenly spaced buckets attached to the circular wheel at one end of the unit the excavated material is fed via a transfer point in the wheel to the belt conveyor system of the excavator for discharge advantage of continuous excavation continuous excavators have reduced dynamic stresses lower service weights reduced maintenance costs and lower power consumption when compared to cyclic excavation tools because of the way they operate and transfer the load bwes have a lower ground bearing pressure allowing more efficient operation in softer underfoot conditions output of continuous excavators continuous excavators are normally rated in terms of their theoretical output where q 60fs swell factor th 10 4 2 where qth is the theoretical output in bank cubic meters per hour cubic yards per hour f is the capacity of each individual bucket s is the number of buckets discharged per hour and the swell factor is that of the material being excavated from table 10 4 1 it is necessary when comparing multiple models or oems to define bucket capacity in the same way in all cases the annual output capability can be calculated by q q oh sf th 10 4 3 where q is the annual output in bank in m3 yd3 qth is the theoretical output calculated in equation 10 4 2 oh is the annual scheduled operating hours and sf is a service factor reflecting operating efficiency and machine availability common service factors range from 0 5 to 0 8 depending on ground conditions climate mana

gerial efficiency and so forth bucket chain versus bucket wheel excavator factors favoring the bce provided selective mining is not required include soft nonabrasive rocks with high deep cuts when the initial box cut is already opened wet pit operations where transport gradients must be reduced where cuts must be taken below grade where specific slope profiles are required and where there are large undulations in the surface of the mineral bed factors favoring the bwe include when selective operation is required where harder ground and or some boulders are encountered when higher availabilities and lower machine maintenance costs are of critical importance when high upward digging capability is needed and where sticky materials are encountered loading equipment selection a number of different types of loading tools are available for use in the mining industry including front end loaders hydraulic excavators and electric cable shovels these are used to directly load material into a haulage unit for transport to a dump or processing facility size selection as with cyclic excavation machines the first step in loading tool selection should be to determine from the mining method planned the planned production rate in whatever units are applicable and then convert these units into a loose volume to be moved per hour equation 10 4 1 can be used to calculate an estimated bucket capacity using the same parameters previously defined theoretical cycle time again values for t can be obtained from information supplied by the oem or from time studies of similar machines in similar conditions in the absence of other data and as a basic assumption to calculate the required bucket capacity a cycle time of 30 to 40 seconds per pass can be used after a specific loading tool is selected a reiteration of this calculation can be performed to calculate a bucket dipper capacity more specifically this range is based on truck positioning on the same level as the loading tool and an 80 to 100 swing if this is not the case allowances should be made to use a slower cycle time bucket fill factor this is a factor of the material sizing condition and the ease or difficulty of filling the bucket table 10 4 1 can be used to approximate the bucket fill factor job operating efficiency as previously defined average operating efficiency figures should range between 0 75 and 0 90 0 83 average mechanical availability availabilities of 85 to 95 should be achievable in most operating conditions during at least the first few years of equipment operation after this calculation is completed one additional adjustment should be made to allow for time lost to exchanging trucks to allow for this lost time if the calculated figure is divided by 0 80 if using a front end loader or single side loading and 0 9 if using a shovel and double side loading alternately loading on either side of the shovel a final bucket capac

ity requirement can be defined the next step in loading tool selection is to consider the type of loading equipment figure 10 4 2 shows the applicable range by bucket capacity for front end loaders hydraulic excavators and electric cable shovels type of loader from the chart in figure 10 4 2 one or two of the loading tool types may be eliminated in this step for example if the required bucket capacity is 15 m3 19 6 yd3 then no electric cable shovels fall into this classification so the choice comes down to a front end loader or a hydraulic excavator likewise if the required bucket capacity is 60 m3 78 5 yd3 then the only choice unless using more than one loading tool to achieve the necessary production is an electric cable shovel table 10 4 3 is a comparison of key points to be considered when selecting among front end loaders hydraulic excavators and electric cable shovels if a hydraulic excavator is selected then a further choice needs to be made between a backhoe front and a loading shovel front a hydraulic front shovel offers many of the same advantages as the electric cable shovel high digging forces good operator visibility along with the additional advantage of better selectivity because the operator can easily see and control where the bucket is placed into the face a backhoe front offers the advantage of less required bench preparation excellent operator visibility into the truck bed to place the load and the ability to reach more material without moving the unit the choice between front shovel or backhoe front is often driven by local preferences most hydraulic excavator manufacturers also offer machines particularly the larger size machines with electric power rather than diesel engine power this allows the mine to take advantage of less expensive electric costs where applicable but some mobility is also lost productivity estimation after a specific loading tool has been selected the productivity estimate can be recalculated using equation 10 4 1 but rearranged to yield the production if truck loading the calculated production rate should be adjusted by multiplying the result by 0 8 or 0 9 based on operating conditions to allow for truck exchange time as described previously haulage truck selection the three types of haulage trucks used in mining operations include on highway trucks articulated trucks and rigid dump trucks the selection will depend on operating conditions and production required on highway trucks on highway trucks in mining applications have a fairly limited application and are most often used to transport materials such as coal commonly in the eastern united states and construction aggregate sand and gravel crushed stone for further processing and or delivery to a customer s site these trucks may be mine owned or more likely are contracted from a local trucking company the size of such units is limited to a maximum of 18 to 20 t 20 to 22 tons selection i

s normally driven by factors such as availability and cost if leased contracted rather than on productivity or size there is however at least one manufacturer of offhighway style trucks that manufactures up to a 36 t 40 ton truck for use within the mine site they purport lower tire costs and lower fuel consumption resulting in a lower operating cost truck than either articulated or rigid frame trucks depending on operating conditions and longevity of the mine though this type of truck could experience a much different pattern of maintenance costs if a mine is considering a truck of this size class a detailed economic analysis should be undertaken to better understand the economics of the different types of trucks figure 10 4 1 can be used to calculate estimated ownership and operating costs a haul study usually done with the assistance of the various oems will assist in calculating production rates several commercially available haul simulation programs are also available however these programs do not normally include on highway trucks in their fleets articulated trucks like on highway trucks articulated trucks have a limited application in mining articulated trucks have a significant advantage over other haulage equipment in areas with soft underfoot conditions and on distances ranging from 120 to 1 200 m 400 to 4 000 ft in longer distance applications and with solid haul road construction rigid dump trucks are the most likely common choice articulated trucks can also be used on steeper grades than rigid dump trucks most oems manufacture articulated trucks with capacities from 22 7 to 36 3 t 25 to 40 tons with a few now manufacturing 45 4 t 50 ton trucks articulated trucks are most commonly loaded by construction class hydraulic excavators with bucket capacities ranging from 2 3 to 4 6 m3 3 0 to 6 0 yd3 the excavator size selection would be made in accordance with the production requirements as shown previously in equation 10 4 1 at that point an acceptable guideline is that a truck should be loaded in four to seven passes table 10 4 4 provides a pass match chart assuming various sized articulated trucks are loaded by various sized hydraulic excavators the number of trucks required can then be calculated by using the number of passes and the pass cycle time to calculate a load time adding fixed times such as positioning at the loading tool and dump time as estimated from experience or from information obtained from the oem and finally adding estimated travel times as calculated using either a haul simulation program or the rimpull speed gradeability curves and retarder charts available from the oem rolling resistance a factor related to how far the truck tires penetrate the haul road surface must also be considered in this calculation when using the charts indicated rigid dump trucks rigid dump trucks are the backbone of haulage equipment for the worldwide mining industry available primarily from

five global oems caterpillar komatsu bucyrus hitachi and liebherr along with other locally manufactured and supported units capacities range from 36 to 360 t 40 to 400 tons with the smaller trucks of 36 to 90 t 40 to 100 tons considered as construction quarry trucks also used in smaller metal nonmetal mining operations and the larger trucks greater than 90 t 100 tons considered as mining class haul trucks size selection mine operators typically target to load a haul truck in three to five passes from the loading tool with hydraulic excavators and front end loaders the range is four to seven passes with electric cable shovels the range is three to four if a truck is loaded in fewer passes the large amount of material dumped at one time can cause excess stress on tires and structural components of the truck if it takes more passes to load the truck waits excessively in the loading area producing inefficiencies table 10 4 5 shows the number of passes required to load different haul truck sizes with various loading tool bucket sizes ranging from 15 to 45 m3 19 5 to 58 9 yd3 major electric cable shovel manufacturers have started to rate buckets by the amount of material carried in each i e 90 t 100 ton load the bucket capacity is sized to match the mine s material density making it an easy calculation to determine the number of passes required to load a truck mechanical versus electric drive one of the primary decisions to be made in haul truck selection is whether to use a mechanical drive truck or an electric drive truck some oems offer both mechanical and electric drive trucks in the mining class greater than 136 t 150 tons while others offer only electric drive within the electric drive option there is also a choice between direct current dc and alternating current ac although ac drive offers additional advantages over dc drive and is becoming the more common choice in today s market electric drive trucks typically travel at higher speed on grades ranging from 4 to 10 have potentially lower maintenance costs offer slightly better fuel economy have a smoother operator ride and offer better retarding capacity to stop the truck however they do have a higher capital cost and require more specialized technical training and capabilities most oems also offer the ability to convert the electric drive truck to a trolleyassist truck resulting in much higher speeds on steep grades mechanical drive trucks can more effectively travel on steeper grades greater than 10 have a larger market presence resulting in more knowledge in the field have a lower capital cost and require a lower level of technical specialization they are also lighter weight vehicles productivity estimation after a loading tool and truck capacity are selected the productivity of the system should be calculated in order to determine the number of trucks necessary and to ensure that the estimated productivity of these d

iscrete units operating together will still meet the required production by far the most efficient way to calculate this productivity is to utilize either one of several commercially available simulation programs or to utilize an equipment manufacturer s program these programs will estimate a production based on the unit load time fixed times for the loading tool and the truck and other data input at least one haul road profile will be required so that the simulation can be performed based on a specific site condition alternatively the only way to estimate fleet productivity is to calculate travel speeds for a given grade using an oem s rimpull speedgradeability chart and add in the applicable fixed times for the loading tool and truck use of computer programs will also allow much quicker evaluations of different loading tools and truck fleet sizes other excavation tools many mining production systems incorporate some combination of excavation and transportation of material these include wheel loaders used in a load and carry application tractor scrapers bulldozers loading draglines and surface miners these are often auxiliary production systems or used in applications with lower production requirements load and carry wheel loaders in addition to truck loading applications wheel loaders can also be used in load and carry applications either to dump directly into an adjacent mined out area such as in an overburden stripping application or to dump into a portable crusher and conveying system the key here is to minimize the distance the loader must travel as longer distances result in slower cycles and increased tire wear a significant cost factor in load and carry operations load and carry applications should be limited to less than 120 m 400 ft productivity of a load and carry operation can be calculated as follows tc time to load bucket travel loaded dump time travel empty 10 4 4 where tc is the loader cycle time in seconds travel time is calculated by dividing the distance to be traversed meters by the travel speed meters per second travel speeds can be determined from oem literature and specifications p qt 3 600 c h 10 4 5 where p is the productivity in cubic meters per hour q is the bucket capacity in cubic meters and tc is the cycle time in seconds calculated in equation 10 4 4 tractor scrapers another excavation tool found in some mining applications particularly overburden stripping with haul distances ranging from 120 to 1 200 m 400 to 4 000 ft is the tractor scraper they may also be found in some medium hardness rock applications that fragment well after blasting many models of scrapers are available with most capacities ranging from 15 to 34 m3 20 to 44 yd3 as noted in the caterpillar performance handbook four primary types of scrapers are available singleengine conventional scrapers tandem powered units elevating scrapers and auger scrapers caterpillar 2008

single engine conventional scrapers have the widest range of applications and commonly require the use of a pusher tractor bulldozer to be loaded most effectively and economically conventional scrapers operate most effectively with lower haul grades lower rolling resistances and better floor conditions tandem powered units offer higher tractive effort capabilities meaning they can be used in softer underfoot conditions and in higher rolling resistance applications normally tandem powered units are also pushed by a pusher tractor to assist in loading elevating scrapers are self loading and best used in applications with short to medium haul distances they also do not perform well in adverse grades and with high rolling resistances elevating scrapers do not perform well in sticky materials such as some clays or material containing rock auger scrapers are also self loading units and like elevating scrapers do not work well with sticky or rocky materials the auger scraper does offer improved tire life and excellent ejection characteristics and is suitable for a range of conditions type selection as detailed in a reference guide to mining machine applications caterpillar 2001 there is some overlap between applications of the different types of scrapers but the following can be used as a basic guideline short distances lower grades elevating or auger short distances higher grades tandem powered longer distances moderate grades conventional low rolling resistances short distances elevating or auger moderate rolling resistances longer distances conventional higher rolling resistances variable distances tandem productivity estimation production with a scraper can be estimated using a methodology similar to that used previously with other types of equipment by considering both fixed times time to load and dump the scraper and variable times travel times loaded and empty utilizing oem rimpull speed gradeability charts total cycle time tct fixed time travel loaded travel empty 10 4 6 fixed time for scrapers will include the load time typically 0 5 to 1 0 minutes depending on type of scraper and actual conditions and the time to maneuver and spread or maneuver and dump the material typically 0 6 to 0 7 minutes scraper productivity can then be calculated as 60 p tct e a swell factor scraper heaped capacity fillability 10 4 7 bulldozers bulldozers are not usually considered as primary excavation tools but can be used to supplement the primary excavation tools one particular application is in coal mine overburden stripping where cast blasting is employed and bulldozers are used to push a significant amount of material into an alreadymined adjacent pit excavation with a bulldozer should be limited to less than 100 m 328 ft to calculate the production that can be achieved by a bulldozer the blade capacity of the bulldozer under consideration can be obtained from the oem literatu

re this would provide a volume per pass in loose cubic meters and can be converted to bank cubic meters by dividing the blade capacity by the swell factor from table 10 4 1 the cycle time can be calculated as t dozing speed distance reverse speed distance c 10 4 8 generally dozing speeds of 1 5 to 2 5 km h 0 9 to 1 5 mph will be the most economical reverse speeds generally in third gear can be found in oem literature when considering bulldozer production the following should be noted if a ripper is used the production when dozing will be increased but the time required to rip will need to be accounted for any available slope should be taken advantage of to doze downhill steady dozing pressure should be maintained speeds should be reduced in areas of heavy shock and impact conditions loading draglines crawler mounted draglines used to load directly into haul trucks have limited application in modern mining methods for many reasons 1 they have limited crowd and breakout action compared to shovels backhoes and front end loaders 2 cycle times are longer when compared to other loading tools resulting in lower production rates and 3 they have limited ability to spot the bucket and dump into a truck several areas in which loading draglines do find an application are in deeper wet pits such as underwater sand and gravel operations and in secondary roles such as excavating box cuts digging sumps final pit cleanup and other areas where they are not the primary production tools estimating production and operating costs of loading draglines can be completed using previously introduced methods adapted to this type of equipment surface miners relatively new to use in surface mining excavation surface miners can be used to excavate material where drilling and blasting is prohibited such as in some quarry applications or where selective mining is required or where materials have a relatively lower compressive strength surface miners can be used in either a discontinuous system where the surface miner cuts the material and loads it into a haul truck or in a continuous system where the material is wind rowed behind or to the side of the machine for later loading usually by a front end loader into a haul truck combinations of a surface miner with a conveyor system can also be achieved in a discontinuous system minimum operating cost is achieved where the truck body is sized so that the truck remains still while the surface miner moves forward and loads the truck uniformly over the full length the maximum volume of each cut can be calculated as 0 6 bucket wheel diameter width of the bucket wheel head 10 4 9 in a discontinuous system the overall productivity will be limited by the number of haul trucks to be used and the distance and haul profile over which the material can be moved according to information from wirtgen 2000 one of the leading surface miner oems in continuous cutting operat

ions the surface miner can achieve outputs up to 1 400 m3 h 1 831 yd3 h bank depending on material compressive strength maximum productivity is achieved in materials with a compressive strength up to 40 mpa 6 000 psi reduced productivity is achieved in materials up to 80 mpa 12 000 psi and in special cases such as small lenses of material or thin layers of up to 120 mpa 18 000 psi some information sources have shown a range of five to seven times higher productivity in material with a compressive strength of 10 mpa 1 500 psi versus a material with a compressive strength of 80 mpa 12 000 psi to establish and maintain competitiveness in international markets for mineral coal and stone products it is necessary to adopt the latest proven technology and economic systems in open cast mining in today s markets overburden of increasing thickness has to be stripped transported and dumped distances to the stockpile are becoming longer depths of mines and quarries are increasing ore grade is decreasing and costs for energy and labor are continuously escalating trucking of waste and ore from mines quarries and pits is a flexible materials handling transportation system mine planners especially at the start of a greenfield project find that trucking is the easiest transportation system to design and plan for as the pit or quarry becomes deeper or farther away from the delivery points mine planners and designers should perform trade offs between cost and flexibility of transportation systems this will ensure that their operations will continue to have the best and most economic materials handling systems for their operations as pits have become deeper and their capacities increased in pit crushing and conveying ipcc has become the comminution and transportation method of choice for most mine planners in response to the ipcc option truck manufacturers developed larger trucks when long term planning is possible ipcc is preferred for the materials handling transportation system there are three main steps in designing an excellent crushing plant 1 process design 2 equipment selection and 3 layout the first two are dictated simply by production requirements and material characteristics but the layout can reflect the inputs preferences and experience of a large number of parties these can include the owner s engineering staff operations and maintenance personnel equipment manufacturers and especially the mine planners the types of in pit crushers usually reviewed by mine planners for hard ore are fixed plants mounted at the rim of the pit and semimobile and fully mobile plants within the pit the gyratory crusher is the crusher of choice for capacities over 2 500 t h metric tons per hour 2 755 stph or short tons per hour for soft rock applications including overburden coal and oil sands fully mobile continuous mining systems are becoming increasingly attractive the crushers most often se

lected for these applications are low speed sizers and double roll crushers two types of conveyor systems being selected for these systems are the conventional conveyor and high angle conveyor the high angle conveyor system has yet to be put in operation in a high tonnage mining operation advantages of ipcc the main reason for the implementation of semimobile and mobile crushing plants instead of fixed crushing plants is the optimization of material transport around and out of the pit on its way to the waste dump or processing plant in this case optimization means the overall cost comparison of truck and conveyor transport in combination with crushing plants stockpiles and dumping equipment the basic design of crushing plants has not changed much in the last few years only the environmental protection requirements for mining companies to install dust suppression and or dust collection equipment have been added almost all ipcc gyratory crusher plants built in the past years are of the semimobile direct dump design the selection and design of a crushing plant depend on evaluation and consideration of the following kind of material to be crushed tonnage of material to be conveyed area depth and development of the open pit space availability height area at the favorable crusher locations taking into consideration the mine design and especially the ore that becomes inaccessible underneath the installation conveying options out of the pit type of downstream material flow in connection with the utilization of the entire crushing and conveying system comparison of maintenance for feeding the crushing plant by direct dump or with an apron feeder advantages of belt conveyor haulage as compared to truck haulage include the following stationing the crusher in the pit reduces cost by shortening the haulage distance between the loaders and crushing plant operating costs associated with fuel tires and lubricants are reduced these products tend to increase in price at a rate that exceeds the rate of monetary inflation labor costs are reduced although most in pit systems either operating or planned use truck haulage the haulage distance is shorter and the number of trucks can be partially or totally reduced this reduction produces a corresponding decrease in operators and maintenance personnel compared with truck haulage safety risks are reduced because mining ventures are long term in pit crushers and conveyors offer greater predictability for future costs dependence on the availability of fuel is greatly reduced dependence on rubber tires is greatly reduced conveyors can traverse grades of up to 30 versus approximately 10 12 for trucks this ability facilitates shorter haulage distances and reduces haulage road construction conveyors can easily cross roads railways waterways and other obstructions with the reduction of haulage costs lower grade ore bodies can be mined economically

this is particularly important because many established ore bodies are decreasing in grade with an increase in depth co2 emissions are greatly reduced downhill conveyors can produce regenerative electrical power instead of dissipating heat as is the situation with trucks braking conveyors are more energy efficient than trucks conveyors require less skilled labor for maintenance than trucks ipcc equipment can achieve maximum operational availability because of greater independence from weather conditions such as fog rain snow and frost the cost of haulage road maintenance is significantly reduced by using conveyors continuous flow of material can be maintained by using conveyors with the availability of technologies such as finite element analysis and computer simulation in pit crushing stations have been refined to a point where their performance and integrity is equal to that of traditional crusher stations by allowing for future relocation as the mine expands long term mine planning is more flexible disadvantages of ipcc the principal drawbacks of belt conveyor haulage as compared to truck haulage are the following short term flexibility is reduced the great mobility of trucks allows mine managers extreme flexibility in the mine plan once an overland conveyor is installed it is prohibitively expensive to move as part of a mine plan change upfront costs are higher while the semimobile system is being moved truck capacity may be insufficient to feed the process or to strip waste remote sites are particularly susceptible to parts shortages truck haulage offers a range of capacities while conveyor systems offer no capacity if parts are not available capacity increments are easier to achieve with trucks compared to large ipcc systems in mines where ore blending is important truck flexibility provides an added advantage lump size is limited once blasted ore and waste in hardrock mines can be loaded directly into a truck and hauled out of the pit generally for conveyors it is necessary to crush the blasted ore or waste a key issue with the ipcc concept has been the inability to maintain projected production over extended periods and maintenance problems since the shutdown of any one piece of equipment in an ipcc system would shut down the whole system the production of the ipcc system is dependent on how it is loaded if the excavator shovel is not working in harmony with the rest of the ipcc system maximum production cannot be obtained on a consistent basis maintenance of the complete ipcc system has to be monitored on a continuous basis the purpose is to minimize the downtime potential of the entire ipcc system while any one individual component is being maintained design parameters the principal design parameters that drive ipcc selection and configuration include production requirements truck sizes capital and operating costs ore characteristics ore body ge

ometry reserve life estimating infrastructure and equipment availability of power and diesel country risks safety and environment project location climate geography terrain life of mine expansion plans operational considerations and maintenance requirements production requirements the process design criteria define the project s production requirements typical requirements are shown in table 10 5 1 crusher station receiving hoppers have to be designed to handle as quickly as possible ore that is delivered by the largest trucks of loading equipment in the fleet the typical design capacity of the feed hopper is two truckloads and in some cases up to three trucks capacity the discharge chamber below the crusher has typically been designed to hold a minimum of 1 25 the capacity of the receiving hopper to prevent the crushed ore from backing up into and damaging the crusher as the standard size of today s mine trucks are 300 to 400 ton capacity the size of the crushing stations must be very large and high just to contain the ore some of the semimobile crushing stations now being designed are direct dump stations with high speed crushed ore receiving conveyors capable of removing the material at rates in excess of the crusher s maximum capacity the crushed rock is then taken directly to the out of pit transport conveyor or into an in pit surge bin use of the high speed belts has reduced the discharge chamber capacity by as much as 50 the initial semimobile gyratory crushers were fed with an inclined apron feeder which allowed the overall height of a station to be contained within two bench heights and allowed for instantaneous dumping of material into the hopper capital costs large semimobile primary crushing plants can be very costly especially if they include the inclined apron feeder it is unwise to estimate crusher installation costs based simply on equipment price plus a contingency allowance for other costs the following direct costs including installation labor hours must all be taken into account earthworks and civil engineering in pit construction planning in order to prevent the interruption of ongoing operations concrete structural steel architectural mechanical electrical and instrumentation indirect costs can be at least half as much again as direct costs and include engineering procurement and construction management start up and commissioning construction equipment spare parts freight taxes escalation and owner s costs relocation hiring and training personnel permits licensing fees etc ore characteristics ore characteristics are a critical element in both crusher and conveyor selection dry ores require greater provisions for dust suppression and collection wet sticky ores can plug chutes and crushers reduce surge capacity and misalign belts for mines at which ore characteristics change over time it can be costly to initially de

sign a plant without the necessary flexibility some owners stipulate that initial capital investment be kept to a minimum with design modifications paid for out of the operating budget this is not always easy to achieve project location a project s geographical location topography remoteness and climate all affect the crusher plant design construction costs are generally much greater at high altitudes in cold climates and at remote sites modular construction and subsequent transportation to the site can improve the economics of such a project geography dictates what material can be best used economically in a particular region a flat quarry operation lends itself to having the conveyor installed in one position for long periods of time a deep copper pit will sometimes require that the crushing station and receiving conveyor need to be moved naturally it would be best to find a wall that requires no more setbacks the conveyor could then be installed either up this face with a high angle conveyor or in a slot designed to install a conventional conveyor another alternative would be similar to the setup at the island copper installation on vancouver island in british columbia canada where the operator installed a conveyor in a tunnel up at 15 out of the pit life of mine expansion plans the life of the mine is a key element in the design of any crushing plant the selection of a fixed crusher versus a semimobile plant is an important design consideration in the overall life of a mine moving a crushing plant and adding feed conveyors to the takeaway conveyor can be expensive any expansion plans for most ipcc systems should be built into the crusher and conveyor systems at the start of a project a conveyor system s tonnage can be increased in the future simply by speeding up the conveyor and if required adding additional drives operational considerations it is important to provide a comfortable well ventilated workspace with drinking water and restroom facilities nearby also the operator should be able to see all the main parts of the crushing facility under his control either through a good window or by means of tv cameras monitors vibration and noise at any crusher station must be kept to a minimum the conveyor should have vehicle access along at least one side maintenance requirements keeping maintenance requirements to a minimum helps achieve higher overall operating availability scheduled preventive maintenance at the crusher station and conveyor involves a number of elements including crusher wear parts feeder wear parts conveyor skirting and adjustment oil and lubrication conveyor belt repair electrical and instrumentation adjustments and visual inspections provisions must be made for either jib or mobile cranes to remove and replace crusher wear parts concaves and the main shaft trolleys jib cranes and pull points should be designed to facilitate equipment maintenance oil and lubric

ation systems should be centralized and designed for easy automatic changes with provisions for well ventilated centralized lubrication rooms where possible process design criteria typically the information required to develop ipcc system design criteria include geographic data climatic data process design data process description ore characteristics etc civil design criteria structural design criteria mechanical design criteria and electrical instrumentation design criteria plant layout and design a carefully designed layout can save significant investment dollars because structures and infrastructure rather than major equipment items represent the major cost element of the crushing plant figure 10 5 1 the mine planner and plant designer must prepare a layout that meets the needs of the design criteria and flow sheet as well as selecting the equipment in the most economic possible configuration it is important to keep structural costs down to design for ease of maintenance and operation and to combine best practices with advances in fabrication and erection most in pit crushing plants are designed by crusher manufacturers so it is imperative that the designer works closely with the selected equipment supplier the manufacturer must remember that the production process economic safety and operational design needs come first three dimensional 3 d computer aided design cad systems tied up with the mine planning 3 d modeling assists greatly in being able to visualize the finished and phased effects of any ipcc installed materials handling transportation system primary crusher selection the crusher is the key to success with any ipcc system the in pit crushing plant can be provided with almost any type of primary rock crusher selection of the primary rock crusher is based on three fundamental considerations 1 type and characteristics of the ore which determine the type of crusher required 2 plant capacity which determines the size of the crusher 3 plant layout and design as the term primary implies primary crushers are used in the first stage of any size reduction cycle the gyratory crusher is the workhorse of the hard rock crushing industry primary gyratory crushers are capable of taking blasted rom and run of quarry feed in size up to 1 500 mm 60 in and producing products ranging in size from 50 to 300 mm 2 to 12 in this type of crusher can sustain production at rates between 318 to 9 072 t h 350 to 10 000 stph depending on the feed characteristics crusher setting and crusher size of the application the primary gyratory crusher is only one of a family of primary crushers that include the following gyratory crusher gyratory jaw double toggle dt jaw crusher single toggle st jaw crusher high speed double roll low speed sizer hybrid roll sizer impactor single shaft hammermill double shaft hammermill feeder breaker all primary crushers can be used

in mobile plants impactors and hammermills are compact and generate a high reduction ratio the high speed of the machines requires special attention to dynamic forces the jaw crushers are good for small tonnage plants and the st jaw crusher has the advantage of being lighter in weight than the dt jaw crusher large capacity jaw crushers result in large crushing plants double roll crushers are large machines in which the two rolls rotate inward so the out of balance forces are minimized these machines are limited to relatively soft and nonabrasive materials however since the return on investment for crushing and conveying systems in the mining industry is heavily dependent on both high capacity and the ability to handle hard and abrasive ores the gyratory crusher has been the reducer of choice throughout the evolution of ipcc table 10 5 2 primary crusher types for large capacity gyratory crusher primary gyratory crushers figure 10 5 2 are typically furnished with radial feed openings of 1 065 mm 42 in 1 370 mm 54 in and 1 500 mm 60 in the largest radial feed opening of any primary gyratory crusher operating in the world is a 1 800 mm 72 in traylor crusher the capacity of even the smallest standard unit the 1 065 mm 42 in gyratory crusher can be sustained at about 2 500 t h 2 572 stph the 1 500 mm 60 in gyratory crusher can crush up to 10 000 t h 11 000 stph depending on the crusher design ore characteristics and desired product size the first 1 500 mm 60 in gyratory crusher was manufactured by the traylor engineering manufacturing company in 1919 at that time the largest haulage trucks available had a 34 t 35 ton payload and shovels were manufactured to match in 2009 the operator had at his disposal 90 yd3 shovels and haulage trucks with a 363 t 400 ton payload truck manufacturers have advised that 500 ton trucks are a possibility tires are the only limitation the result of larger haulage trucks is an obvious mismatch between the top size of ore fed to the crusher and the largest radial feed opening available the consequence of this mismatch is the bridging of two or more large lumps that have been fed to the crusher at the same time bridging has been partially compensated for by the use of hydraulic rock breakers installed on pedestal mounted booms in some installations the hydraulic rock breaker is employed up to one fifth of the total time the crusher is in operation with 20 spent breaking oversize rock and 80 spent breaking bridges bridging has also been mitigated by improvements in blasting technology the largest standard size gyratory crusher the 1 500 2 970 mm 60 117 in typically operates at the upper limit of the capacity range these machines have been optimized with improvements in gearing bushing materials and lubrication systems the speed has been increased greater horsepower is available from the manufacturers the crushing chambers have been optimiz

ed with cad larger size crushers are available and the next size will most likely be a 1 800 3 250 mm 72 128 in however the increased capacity of between 7 and 10 does not seem to provide a cost effective benefit because of the increase in weight height and cost for both the crusher and the auxiliary equipment to support the crusher double roll crushers in the past few years the high speed double roll crusher figure 10 5 3 has been gaining respect in many circles primarily oil sands and overburden the capacity of the double roll crusher can exceed 14 000 t h 15 432 stph because of its size and method of processing the ore at this capacity the product size is generally 400 mm 16 in which is acceptable for oil sands because secondary processing and ablation during slurry transport will reduce the lump to a mixture of sand bitumen clay and water this degree of size reduction is also acceptable for waste handling as the ore needs to be crushed to a size suitable for conveyor transport the double roll crusher is suitable for ore that is wet sticky compactable or has a high silica content indeed every set of properties that causes problems for just about every other type of crusher the wear elements are the crusher teeth which can be easily replaced in fact the double roll crusher operates best with a mixture of new and worn teeth so a worn crusher is still a performing crusher the double roll crusher is capable of processing massive lumps which in oil sands can exceed 7 m 23 ft in one dimension the shovel operators try to mine selectively but in the dead of winter the steam rising from the working face usually obscures the operator s view of the face especially at night and as a result selective mining is not very accurate low speed sizers in the early 1980s low speed sizers were introduced figure 10 5 4 this represented one of the only fundamental developments in primary crushers in three quarters of a century the main technical feature of the low speed sizer which can broadly be considered a variety of toothed roll is that it exploits the fact that the ratio of compressive strength to tensile and shear strength in the majority of rocks is approximately 10 1 the low speed sizer breaks the rock in tension or in shear by snapping and chopping action rather than in compression as conventional crushers do additionally the positioning of the teeth on the rolls allows undersize to fall directly through the machine resulting in high throughputs at very low rotational speeds which means greatly reduced wear energy savings better control of discharge size in three dimensions and greatly reduced fines low speed sizers are used for soft to medium hard nonabrasive sticky types of materials up to 200 mpa 29 000 psi for example coal oil sands medium hard limestone kimberlite gypsum clay shale schist and gold ore these sizers are also used to crush bauxite and overburden wher

e the host rock is relatively soft and the inclusions range up to 400 mpa 60 000 psi the low speed sizer is not particularly sensitive to abrasion if the reduction ratio is low low speed sizers can be fabricated so that the frame can accommodate material larger than 2 000 mm 70 in as truck capacities continue to grow and rom feed continues to increase hybrid roll sizer the hybrid roll sizer figure 10 5 5 features a compact design as a result of the roll diameter made possible by the aggressive tooth geometry thus minimum space requirements and the distinctive advantages of a classic double roll crusher such as hydraulic gap adjustment and overload protection are combined in one machine furthermore the hybrid is able to compensate peaks with energy storage in the flywheels the advantages of a hybrid roll sizer are compact design minimum generation of unusable fine material high throughput capacities up to 11 000 t h 12 125 stph hydraulic gap adjustment system overload protection and processing of wet and sticky materials the disadvantages are unsuitable for very hard and extremely hard materials low reduction ratio and not economic for low tonnages unless the material is difficult to handle types of in pit crushing systems the in pit crushing systems developed and operated to date have varying degrees of mobility ranging from fully mobile units to permanently fixed plants which resemble traditional in ground crushing plants the crushing plants can be stationary mounted on concrete foundations or semimobile style supported on steel pontoon feet as the mining operation progresses the semimobile crushing plants can be relocated within the mine using multiwheeled trailers or transport crawlers typically shovels load the rom ore on to heavyduty haul trucks that transport the ore to the crushing plant thus relocating the crushing plant as the mine expands and reducing the distance that the large trucks need to haul the ore from the working face the following terms are presented to help distinguish the range of mobility within the generic term of in pit crushing systems fixed crushers stationary in ground or rim mounted crushing plants the stationary in ground crusher is installed in a concrete structure below grade the crusher is usually located external to the pit and is never moved the stationary rim mounted crusher is typically installed in a concrete structure which is part of or attached to the bench wing wall figure 10 5 6 a portion or all of the structure may be fabricated steel and could be disassembled and moved figure 10 5 7 the stationary rim mounted crusher is usually installed for 15 or more years semifixed crushing plant the semifixed crusher is mounted on a steel structure that rests on a concrete foundation figure 10 5 8 the structure houses some or all of the auxiliary equipment and subsystems to operate the crusher the crusher is located at or near the edge of the

pit some degree of disassembly is required to move the structure the planned frequency of moves for a semifixed crusher is no less than 5 to 10 years semimobile indirect feed crushing plant the semimobile indirect feed crushing plant is an all steel structure figure 10 5 9 the plant typically consists of three major modules the apron feeder the crushing plant with the crusher and a separate tower that houses the control room the control room module is bolted to the crusher module when the plant is moved the civil work required for retaining walls is relatively simple and offsets the cost of the apron feeder however the work is massive because the structures often in excess of 25 m 82 ft in height are required to support the load of the 400 ton trucks the crusher is typically located near the centroid of the working portion of the mine to minimize truck haul distance to allow for movement of the structure by commercially available transport equipment bulkheads are built into the structure the planned frequency of moves for a movable crusher is between 3 and 5 years semimobile direct dump crushing plant the semimobile direct dump crushing plant is mounted on a steel structure that houses all of the auxiliary equipment and subsystems to operate the crusher figure 10 5 10 the structure is self supporting and rests on the mine floor either with or without footers the plant design allows for two or three dump points to minimize truck haul distance the crusher is typically located near the centroid of the working portion of the mine bulkheads are built into the structure to allow for movement of the structure by commercially available transport equipment the planned frequency of moves for a movable crusher is between 1 and 10 years fully mobile crushing plant the fully mobile crusher is mounted on a steel platform and is self propelled figure 10 5 11 the platform houses all auxiliary equipment and subsystems to operate the crusher and is self supported and rests on the mine floor to minimize truck or front end loader haulage the crusher is located at the working face wheels crawlers or pneumatic pads are integrated into the platform and drive power to move the equipment is included on board the planned frequency of moves for a fully mobile crusher is between 1 day and 1 week fully mobile continuous crushing system the fully mobile in pit continuous crushing system is mounted on a steel platform and is self propelled figures 10 5 12 and 10 5 13 the platform houses the apron feeder crusher discharge conveyor and all auxiliary equipment and subsystems to operate and propel the crushing plant the feed hopper needs to accept rom material that is dumped from shovels and or draglines the platform is self supported and rests on the mine floor the crusher is located at the working face for direct feed crawlers are integrated into the platform and drive power to move the equipment is included on board the pl

ant moves in tandem with the shovel the conveying system moves in tandem with the in pit continuous crushing system operation of ipcc crushing plants fixed rim mounted crushers fixed rim mounted crushers are increasingly incorporating traditional direct dump arrangements with these designs the hopper above the gyratory crusher is designed to hold 1 5 to 2 times the capacity of the largest truck that will dump into the crusher during operation discharge surge bins hav traditionally been sized slightly larger than the feed hopper to accommodate any unusual fines condition in order to reduce overall height and thus capital costs discharge apron feeders have been replaced by impact resistant discharge belt conveyors the trend away from discharge apron feeders to discharge belt conveyors has allowed for wider belts with greater capacity in conjunction with high capacity discharge belt conveyors the typical capacity of the surge bin below the crusher has decreased dramatically even with removing the discharge apron feeder and reducing surge bin capacity direct dump arrangements result in tall structures with rim mounted in pit crushers this tall overall height requires wing walls to support and reinforce the structure traditional fixed crushers are installed below grade and fed at grade recent in pit installations such as at the cripple creek and victor gold mine in colorado united states which is an all concrete structure and buxton lime in buxton england which is an all steel structure are partially below and above grade to accommodate a single bench height the dump pockets can be arranged for one two or three dump positions with a two position dump pocket design the two dump points are set 90 apart from each other the spider orientation is either in line with the centerline between the two dump positions or at 90 to the centerline between the two dump positions either position is mechanically satisfactory to the gyratory crusher with a three position dump pocket design it is universally accepted that the spider is orientated in line with the centerline of the central dump position advantages of fixed rim mounted crushers with direct feed arrangements include traditional plants with simple configurations easily adapted for in pit crushers reduced maintenance costs due to no longer needing an apron feeder high crushing chamber throughput reduced capital costs due to limited degree of mobility reduced maintenance costs due to a greater amount of crushing in the upper portion of the chamber and decreased localized abrasive wear and greater capacity and finer product size due to the weight of the ore column above the crusher disadvantages of fixed rim mounted crushers include the following poured concrete design cannot be moved structural steel designs are typically not designed to be moved if the structure were designed to be moved an extensive substructure is required to support the plant fo

r moving overall height is greater because of the higher dump point bench level greater height means extensive retaining wall structures semifixed crushers semifixed crushers are the mining industry s attempt to incorporate the advantages of limited mobility while eliminating the need for an extremely expensive inclined apron feeder semifixed plants have incorporated both indirect feed using a horizontal apron feeder as well as various forms of directdump arrangements due to the high capital and maintenance costs of apron feeders as well as the availability of high payload haulage trucks capable of sustaining high crushing capacities the majority of semifixed crushing plants installed since about 2000 have incorporated direct dump arrangements in a semifixed crushing plant a portion of the station is fabricated from steel the direct dump feed hopper crusher support structure and control rooms are almost always steel fabricated differences in design are related to the degree where the lower portion of the plant is concrete or steel typically only the crusher and part or all of the dump hopper are mounted on a steel base while the remainder of the station is of civil construction the steel portion is moved by crawler transporter as self propelled modular transporters spmts the station is moved to a new civil structure and the old station destroyed advantages of semifixed crushers with direct feed arrangements include traditional plants with simple configurations easily adapted for in pit crushers reduced maintenance costs due to no longer needing an apron feeder high crushing chamber throughput reduced capital costs due to limited degree of mobility increased long term flexibility due to the limited mobility which allows for future changes and modifications reduced maintenance cost due to greater amount of crushing in the upper portion of the chamber and decreased localized abrasive wear and greater capacity and finer product size due to the weight of the ore column disadvantages of semifixed crushers with either indirect or direct feed arrangements include only the crusher and part or all of the hopper are mounted on a steel base and the balance of the station is civil construction greater overall height is due to the higher dump point bench level semimobile indirect feed crushing station the semimobile indirect feed in pit crushing stations utilize an apron feeder to lift ore into the feed opening of the primary crusher the use of an apron feeder allows for the crusher station to either operate at grade or to utilize a single low bench the semimobile indirect feed in pit crushing station is typically designed and built in three or more modules apron feeder primary crusher and the control room some semimobile controlled feed in pit crushing stations have a separate lubrication and hydraulic system module located adjacent to the main structure new environmental controls for dust collecti

on and or dust suppression may add a dust collection module as a separate item the control room is separate from the crusher structure to reduce crusher initiated vibration in the control and electrical rooms the design usually incorporates a feature that allows for the control room to be attached to the crusher module for moving additionally the use of a truck dump hopper at the apron feeder creates a large surge pocket between the mine and the crusher making the flow of ore through the crusher more uniform and continuous advantages of indirect feed using an apron feeder include low bench height for dumping ore reduced truck queue time due to the surge pocket improved control of oversize material fed to the crusher and reduced crusher downtime due to bridging of large lumps disadvantages of indirect feed using an apron feeder include increased total capital cost increased maintenance costs associated with adding an apron feeder and increased maintenance costs associated with the crusher from using an apron feeder due to the nature of the feeder ore tends to impinge upon small areas within the crushing chamber causing premature localized wear of the concaves and mantles the use of alloy steels has mitigated the problem although the cost of alloy steel components remains higher than manganese steel and availability is limited semimobile direct dump crushing plant the semimobile direct dump crushing plant has been the design of choice for ipcc since about 2000 this design incorporates all of the features of the traditional in ground crushing station the crushing plant incorporates the feed hopper the crusher and the lubrication and hydraulic systems to support the crusher as well as all maintenance equipment including a rock breaker and usually a crane with capacity to lift the mainshaft assembly with the oversize mantle an operator s control room and electrical rooms are also included above the traditional gyratory crusher station with two dump points and an apron feeder discharge is the hopper which can hold two times the capacity of the largest truck that will dump into the crusher during operation for an operation with 400 ton trucks this translates into 726 t 800 st live ore capacity traditional design for the surge pocket under the crusher of a gyratory crusher plant would be 2 5 times the largest truck or 907 t 1 000 st capacity to prevent backup of ore into the crusher chamber in case of an unusual fines condition in order to reduce overall height and thus capital costs discharge apron feeders have been replaced by impact resistant higherspeed discharge belt conveyors the trend away from discharge apron feeders to discharge belt conveyors has allowed for wider belts with greater capacity in conjunction with high capacity discharge belt conveyors the typical capacity of the surge bin below the crusher has decreased dramatically because the crusher discharge belt has a higher capacity th

an the haulage belt a surge bin is usually added to the circuit the structure is self supporting and rests on the mine floor either with or without footers the plant design allows for two or three dump points the crusher is typically located near the centroid of the working portion of the mine to minimize truck haul distance bulkheads are built into the structure to allow for movement of the structure by commercially available transport equipment the planned frequency of moves for a movable crusher is between 5 and 10 years advantages of semimobile direct feed arrangements include traditional plant configuration reduced maintenance costs due to deletion of the apron feeder high crushing chamber throughput reduced capital costs due to limited degree of mobility increased long term flexibility due to the ability to move the complete station intact reduced maintenance costs due to greater amount of crushing in the upper portion of the chamber and decreased localized abrasive wear when compared to indirect feed designs and greater capacity and finer product size due to the weight of the ore column disadvantages of semimobile crushers with either indirect or direct feed arrangements include large and heavy structure requiring large transporters for moving and greater overall height due to the higher dump point bench level which requires extensive bench retaining walls fully mobile crushers the fully mobile crusher is mounted on a steel platform and is self propelled wheels crawlers or pneumatic pads are integrated into the platform to move the station the platform which is self supported and rests on the mine floor houses all auxiliary equipment and subsystems to operate the crusher to minimize truck or front end loader haulage the crusher is located at the working face the planned frequency of moves for a fully mobile crusher is between 1 day and 1 week the majority of fully mobile in pit crushing stations utilize an apron feeder to lift ore into the feed opening of the primary crusher the use of an apron feeder allows for the crusher station to operate at grade and move without the need for employing spmts for hard rock including aggregate and cement the capacity has been limited to about 2 500 t h 2 756 stph with gyratory crushers the 1 370 1 880 mm 54 74 in gyratory crusher is the largest unit used for this application mobile crushing plants are favorably installed under the following conditions clear and undisturbed geological situation almost even and horizontal coal and waste rock layers straight benches as long as possible for shiftable face conveyor installation long term mine planning for the design of face conveyors collection conveyors at bench end side slopes and connecting conveyors to waste dumps or processing plants advantages of fully mobile crushers include elimination of truck transport reduced number of personnel avoidance of high truck maintenance cos

ts reduction of mine traffic and increase in overall safety disadvantages of fully mobile crushers include increased total capital costs increased maintenance costs associated with adding an apron feeder and increased maintenance costs associated with the crusher from using an apron feeder due to the nature of the feeder ore tends to impinge upon small areas within the crushing chamber causing premature localized wear of the concaves and mantles the use of alloy steels has mitigated the problem although the cost of alloy steel components remains higher than manganese steel and availability is limited fully mobile continuous crushing systems the fully mobile continuous crushing system includes the fully mobile crushing plant as one component of the system figure 10 5 14 the components of the system are the shovel or dragline the fully mobile crushing plant and the conveying system that transports the crushed material to the next operation in the flow sheet for fully mobile continuous crushing systems handling overburden the conveyors take the material to a dump for coal and oil sands the material is transported to the processing plant the fully mobile crushing plant takes the material directly from the shovel or dragline the full bucket load is discharged into the hopper of the fully mobile crushing plant an apron feeder elevates the material and discharges it into the crusher either a low speed sizer a high speed double roll crusher or a hybrid roll sizer the product from the crusher is transported to the overland conveyor system all parts of the system are designed to work continuously this requirement has led to the development of fully mobile crushing plants that have the ability to continue operation while moving older systems in oil sands in canada and coal in australia had to stop operations lift the apron feeder off the ground move to a new location and set the apron feeder down before resuming operation mobile crushing plants are able to work at one two or three benches with only one shiftable bench conveyor multibench operation an important design criterion for this effective technology is the transfer equipment to the bench conveyor which has to bridge not only distances in length but also in height crusher discharge conveyor mobile transfer conveyors and mobile transfer bridges ramps in the working face or in the side slope are necessary for equipment relocation from one bench to the other ipcc process control growing recognition exists in the industry of the effect that mining practices have on the efficiency of mineral processing operations among numerous variables size distribution of ore is widely accepted as having a significant effect on the throughput and recovery achieved in mineral processing using available digital imaging technology improved process monitoring and modeling provide the opportunity not only to identify these variables but also to monitor and control

them throughout the entire mining process in real time the majority of mine designs employ crushers to reduce the size of the material originating from the mine the primary crusher is the link between the chemical comminution blasting and the beginning of the mechanical comminution circuit crushing and milling as such an in pit crusher is not only a key point in the process to apply a measurement monitor but also a key resource to be optimized feed to the primary crusher from the mine can be measured and monitored to establish blasting performance size information associated with each haul truck can be traced back to the location in the mine plan and used to help design future blasting practices the crusher product is usually the beginning of the mineral processing circuit that involves more energy consumption to further reduce fragment size in the short term the crusher s performance is the responsibility of the crusher operator who through the use of digital imaging now has a record of the quality of the crusher product size and can make adjustments to the crusher to keep the product in specification as required by the design of the remainder of the comminution circuit in the long term archives of size trends of crusher feed and product as related to other key performance indicators such as blast fragmentation crusher mill throughput crusher reduction ratios bond work index and energy consumption and efficiency optimize managementlevel decisions opportunities for optimization include how to tailor blasting to feed the stationary mechanical comminution circuit how to load the crusher and how to establish better proactive maintenance of the crusher as well as keeping tighter specifications on the feed and products of the various stages of comminution by applying digital imagery technology a technology widely used in numerous other manufacturing industries new innovative solutions are available to the mining industry applying imaging technology can generate volumes of size information which were not previously possible for use in long term studies as well as in short term process control to create long term operating savings as compared to manual sizing methods or worse yet no measurement at all an engineer with more than 30 years of mining experience was quoted about image analysis at our mine we utilize digital image analysis systems to provide quality quantitative fragmentation information on our blasting and integrate the fragmentation information into our operations database as a quality control mechanism within our ongoing continuous improvement program an important step to controlling costs is controlling your process basically any company that has a product that requires control of particle size and is concerned with profitability needs this valuable information introduction in truck based hauling systems the mine haul road network is a critical and vital component of the production process

as such underperformance of a haul road will have an immediate impact on mine productivity and costs operations safety productivity and equipment longevity are all dependent on welldesigned well constructed and well maintained haul roads thompson and visser 1999 the mine haul road is an asset and should in conjunction with the haul trucks using the road be optimally designed and its routine maintenance managed accordingly an ad hoc or empirical approach to haul road design is generally unsatisfactory because it has the potential for overexpenditure both on construction and operating costs arising as the result of the overdesign and over specification of short term low traffic volume roads and the underdesign leading to excessive operating and road maintenance costs and premature failure of longer term higher volume roads economies of scale and the increase in haul truck payload have led the ultra class truck 220 t metric tons and larger population to rise to more than 40 of all mine trucks used gilewicz 2006 with this increasing size haul road performance can be compromised resulting in excessive total roaduser costs translating to an increase in cost per ton hauled but also indirectly to reduced production rates and vehicle and component service life truck haulage costs can account for up to 50 of the total operating costs incurred by a surface mine and any savings generated from improved road design and management benefit the mining company directly as a reduced cost per metric ton of material hauled central to the cost of truck hauling is the concept of rolling resistance expressed here as a percentage of gross vehicle mass gvm it is a measure of the extra resistance to motion that a haul truck experiences and is influenced by tire flexing internal friction and most importantly wheel load and road conditions empirical estimations of rolling resistance based on tire penetration specify typically a 0 6 increase in rolling resistance per centimeter tire penetration into the road over and above the 1 5 radial and dual wheel assemblies to 2 cross ply or single wheel assemblies minimum resistance in addition to tire penetration road surface deflection or flexing will also generate similar results with the truck tire running up grade as the deflection wave pushes ahead of the vehicle taking an electric drive rear dump ultra class truck of 376 t gvm as an example on a ramp road with a basic rolling resistance of 2 an additional 1 rolling resistance will reduce truck speed by 10 to 13 whereas on a flat surface road the truck speed will be reduced from 18 to 26 although many concepts from highway engineering can be adapted to the design construction and management of mine roads significant differences in applied loads traffic volumes construction material quality and availability together with design life and road user cost considerations mitigate for a tailored design solut

ion for mine haul roads components of an integrated mine haul road design the operating performance of a mine road can be subdivided into four design components i e geometric structural functional and maintenance management when designing and constructing a haul road for optimal performance these design components are best addressed using an integrated approach if one design component is deficient the other components may not work to their maximum potential and road performance is often compromised this will most often be seen as maintenance intensive or high rolling resistance roads translating to increased equipment operating downtime and repair costs the cure however is not necessarily just more frequent maintenance no amount of maintenance will fix a poorly designed road design and management of haul road systems should also be approached holistically especially with regard to the benefits achieved from various solutions to enhance productivity although for instance trolley assist may improve cycle times and reduce cost per metric ton hauled it is first necessary to evaluate the extent to which an existing haul road network meets optimal design requirements before resorting to solutions that do not directly address the key deficiencies of the existing road system the recommended approach is therefore to assess the extent to which the asset the current road network exhibits scope for improvement and once optimized then revert to resource supplementation to leverage these benefits through optimal asset and resource interaction figure 10 6 1 illustrates the approach to mine road design based on the geometric structural layerworks functional wearing course and maintenance management design components together with a dust palliative evaluation methodology these design components form the basis for the following sections of the chapter the first component that of geometric design is commonly the starting point for any haul road design and refers to the layout and alignment of the road in both the horizontal and vertical plane the ultimate aim to produce an optimally efficient and safe geometric design can only be achieved when sound geometric design principles are applied in conjunction with the optimal structural functional and maintenance management designs the aim of a structural design is to provide a haul road that can carry the imposed loads over the design life of the road without the need for excessive maintenance it is focused on the design of road layerworks and the response of construction materials in and under the road to the truck wheel loads the functional design is centered on the selection of wearing course or surfacing materials the most suitable choice application technique and maintenance strategy is required commonly the running surface of a mine road is a gravel mix which lends itself to maintenance blading or over the longer term rehabilitation to improve p

erformance of the material palliation and or stabilization is often considered primarily to reduce both dust generation and material degeneration the latter leading to increased rolling resistance and associated road maintenance the maintenance aspect of haul road design cannot be considered separate from the geometric structural and functional design aspects because they are mutually inclusive design and construction costs for the majority of haul roads represent only a small proportion of the total operating and maintenance costs although it is possible to construct a mine haul road that requires no maintenance over its service life construction costs would be prohibitively expensive the converse an empirically designed and cheaply constructed road would also incur excessive costs in this case related to vehicle operating and road and vehicle maintenance costs the use of an appropriate road maintenance management strategy will generate significant cost savings by virtue of a better understanding of the relationship between wearing course material degeneration rates manifest as increasing rolling resistance on the road and its influence on both cost per metric ton hauled and the cost of road maintenance itself a mine road network often comprises various roads each with a specific function traffic type size of truck traffic volume service level performance and operating life a road classification system should be developed according to these parameters as part of a mine wide common framework for road design this can be used as the starting point for design guidelines for construction personnel to enable them to determine easily what design guideline is appropriate when constructing new roads or evaluating and rehabilitating existing mine roads clearly not all roads are equal and thus the approach to design and management must be tailored to apply more resources to high volume long term and highcost impact road segments across the network figure 10 6 2 illustrates typical haul road design categories the accompanying data forms the basic input to the four design categories previously discussed geometric design the geometric layout of a mine haul road is dictated to a great extent by the mining method used and the geometry of both the mining area and the ore body mine planning software enables various haul road geometric options to be considered and the optimal layout selected both from a road design and economic lowest cost of provision perspective minemap 2008 although these techniques often have default design values embedded in the software it is nevertheless necessary to review the basic concepts of geometric design if any modifications are to be considered in the design of mine roads either on the basis of economics or more critically from a safety perspective the road layout or alignment both horizontally and vertically is generally the starting point of the geometric design practically it

is often necessary to compromise between an ideal layout and what mining geometry and economics will allow any departure from the ideal specifications will result in reductions in both road and truck life considerable data already exists pertaining to good engineering practice in geometric design kaufman and ault 1977 usbm 1981 tannant and regensburg 2000 and forms the basis of the design criteria developed here broadly speaking safety and good engineering practice require haul road alignment to be designed to suit all vehicle types using the road operating within the safe performance envelope of the vehicle or where this is not possible at the speed limit applied ideally geometric layout should allow the vehicles to operate at their maximum safe speed but given that the same road is used for laden and unladen haulage there is often the need to minimize laden travel times through appropriate geometric alignment while accepting compromises generally in the form of speed limits on the unladen return haul the process of geometric design is shown in figures 10 6 3 and 10 6 4 and discussed in the following sections vertical alignment vertical alignment considers the effect of road elevation changes on both sight and stopping distances of trucks together with the effect of road grade on safety and performance sight distances at least 150 m is required based on typical stopping distance requirements on a curve or bend in the road this could be difficult to achieve as shown in figure 10 6 5 length of vertical curves l can be determined from consideration of the height of the driver above the ground h1 m an object of height h2 m usually 0 15 m to represent a prostrate figure in the road sd as the minimum stopping distance m and g as the algebraic difference in grades optimal and maximum sustained grades although maximum gradients may be limited by local regulations ideally the gradient should be a smooth even grade not a combination of grades laden trucks running against the grade work best at total i e grade and rolling resistance values of between 8 and 11 however each truck engine and drive system combination has a characteristic optimal grade curve and it is a good geometric design starting point to determine the optimal gradient for the selected truck in use at the mine although travel times laden are sensitive to grades against the load care should also be taken when selecting the grade from the perspective of truck retard limitations on the unladen downward leg of the haul this aspect becomes critical in the case of downgrade laden hauling when retard capacity would be the limiting design criteria horizontal longitudinal alignment haul road width curve layout and associated superelevation and run out together with cross slope or camber and intersection layout are incorporated under longitudinal geometric design considerations width of road the width of the road should allow

enough room for the required number of lanes and all the associated safety and drainage features the widest vehicles proposed determine the roadway width the dimensions of the safety berms and drainage channels are also added to the roadway width to determine the construction width required table 10 6 1 summarizes these design roadway widths on single lane roads if the sight distance is less than the stopping distance sufficient space must be provided for moving vehicles to avoid collision with stalled or slow moving vehicles curvature and switchbacks any curves or switchbacks should be designed with the maximum radius possible and be kept smooth and consistent changes in curves radii compound curves should be avoided a larger curve radius allows a higher safe road speed and increased truck stability sharp curves or switchbacks will increase truck cycle times and haul cost as a result of rear dual tire wear due to tire slip curve superelevation banking this is the amount of banking applied on the outside of a curve to allow the truck to run through the curve at speed ideally the outward centrifugal force experienced by the truck should be balanced by the side friction between tires and road superelevations should not exceed 5 to 7 unless highspeed haulage is maintained and the possibility of sliding minimized table 10 6 2 shows typical superelevation rates based on speed of vehicle and radius of curve where tighter curves are required or truck speed is higher on approach to the curve a speed limit should be applied cross slope or camber this is critical to the design and successful operation of mine roads applying a crossfall or camber ensures that water does not gather on and penetrate into the road surface standing water on or in a road will cause rapid deterioration of the road two options exist either a cross slope from one edge of the road to the other edge or a camber crown from the center of the road to both sides of the road whatever option is adopted at the point where the road edge and camber or crossslope downslopes meet a drainage ditch must be provided the two options are illustrated in figure 10 6 7 with safety berms omitted but including drain locations a constant camber or cross slope of 2 to 3 is ideal providing adequate drainage without incurring adverse truck tire and strut loading a preference may exist for cross slopes because of the envisaged equalized load sharing and tire scrub i e loss of tire tread rubber because of shear contract with the road wearing course a cross slope should be used with caution with the slope falling toward the inside of the bench toe position as opposed to the outslope side where a camber or crown is selected and where this leads to the possibility of trucks sliding in the direction of the bench crest or outslope or toward a large vertical drop it is only recommended when large deflection berms are placed at road crest edge special considera

tion must be given to determining when to use the maximum and minimum rates lower cross slopes are applicable to relatively smooth compact road surfaces that can rapidly dissipate surface water without the water penetrating into the road surface in situations where the surface is relatively rough a larger cross slope is advisable run out run out should be used when a section of haul road changes from a cross slope or camber into and out of a superelevated section the change should be introduced gradually to prevent excessive twisting or racking of the truck chassis the run out length is typically apportioned 25 to 33 to the curve and 66 to 75 to the tangent or run up to the curve intersection layout this often entails smaller curve radii than that required for maintaining high speed hauling junction priorities should be maximized for nonyield traffic flows the design of the intersection run in and run out should be a smooth spiral progression from one cross slope or camber to the curve superelevation and out again as shown in figure 10 6 8 drainage at the side of each intersection leg must be adequate to keep water from ponding at the roadside and where possible intersection segments or at least nonyield segments must be placed on flat not inclined areas where feasible and safe to do so junctions can be designed on a herringbone layout with traffic flow in the direction of the primary destination i e the largest radius of curvature out of the junction which will facilitate higherspeed hauling and reduce tire scrub but consideration should be given to maintaining sight distances in all four quadrants combined alignment when laying out a haul road some additional considerations arise with combined alignment sharp horizontal curves at or near the top of a grade section should be avoided if a horizontal curve is necessary it should start well in advance of the vertical curve switchbacks should be avoided where possible but if mine plan dictates their use the radius should be as large as possible and should not be placed on grade sharp horizontal curves requiring a further speed reduction following a long sustained downgrade where haul trucks are normally at their highest speed should be avoided short tangents and varying grades should be avoided especially on multilane roads grades should be smooth and of consistent grade percentages intersections should be avoided near the crest of vertical curves or sharp horizontal curves intersections should be as flat as possible with sight distances being considered in all four quadrants where an intersection lies at the top of a ramp 100 200 m of flat road should be considered before the intersection and stopping and starting a laden haul truck on grade should be avoided drainage design at intersections should stop any ponding of water against intersection superelevated curves sections of road with no camber or cross fall should be avoided often

encountered at curve superelevation runin or run out these flat sections should preferably be at a 1 to 2 vertical grade to assist drainage ditches and drainage a well designed drainage system is critical for effective operation and safety water on the road or in the road layers will quickly lead to poor road conditions as part of the haul road geometric design process contours in the vicinity of the proposed road should be examined prior to construction to identify areas of potential ponding locations of culverts and so forth drains at the edge of the road should be designed to lead the water off the road without causing erosion drains should not be cut into the base layer and should preferably be lined with compacted wearing course thereby preventing water from seeping into the underlying layers v ditches are recommended for nearly all applications owing to their relative ease of design construction and maintenance the ditch cross slope adjacent to the haul road should be 4h 1v or flatter except in extreme restrictive conditions in no case should it exceed a 2h 1v slope the outside ditch slope will vary with the material encountered in rock it may approach a vertical slope in less consolidated material it may be a 2h 1v slope or flatter in a cut fill section a cross slope toward the cut side should be used and drainage should run in a single ditch in a total cut or total fill section drainage should be carried on both sides with crown or camber from the road center line ditch lining is a function of road grade and in situ material characteristics between 0 and 4 grade the ditch may be constructed without benefit of a liner except in extremely erodible material such as sand or easily weathered shale and silts at grades greater than 5 the lining should consist of coarse crushed waste rock placed evenly on both sides to a height no less than 0 3 m above the maximum depth ditches must be designed to handle expected runoff flows adequately under various slope conditions the primary consideration is the amount of water that will be intercepted by the ditch during a rainstorm typically a 10 year 24 hour storm chart should govern the design culvert sections are used to conduct runoff water from drainage ditches under the haul road if buried piping is used they should be set at 3 to 4 fall and smooth wall concrete pipes should be used in conjunction with a drop box culvert of a size that will allow it to be cleaned with a small backhoe excavator at all culvert inlets a protective encasement or headwall consisting of a stable nonerodible material should be provided typical culvert units are either portal and rectangular precast concrete culvert units or precast concrete pipe culvert units depth of cover after the culvert pipe is determined by the type of culvert in relation to the vehicles that will use the road a minimum cover of 1 000 mm over the pipe is required in most cases all prefabricated culv

erts should be constructed under trenched conditions after the road has been constructed concrete pipe culverts are laid on a layer of fine granular material 75 mm thick after the bottom of the excavation has been shaped to conform to the lower part of the pipe where rock shale or other hard material is encountered on the bottom of excavations culverts should be placed on an equalizing bed of sand or gravel after placement the culvert trench is backfilled and compacted structural design structural design concerns the ability of a haul road to carry the imposed loads without the need for excessive maintenance or rehabilitation haul roads deteriorate with time because of the interactive effort of traffic load and specific subgrade and in situ material strengths and structural thicknesses the california bearing ratio cbr method kaufman and ault 1977 has been widely applied to the design of mine haul roads in which untreated materials are used however when multilayered roads are considered in conjunction with a base layer of selected blasted waste rock a mechanistic approach is more appropriate when a selected waste rock layer is located under the wearing course road performance is significantly improved primarily because of the load carrying capacity of the waste rock layer which reduces the susceptibility of the soft subgrade and in situ material to the effects of high axle loads it also has the added advantage of reduced construction costs by virtue of reduced volumetric and compaction requirements compared with the cbr cover curve design approach morgan et al 1994 thompson and visser 1996 a mechanistic design is based on a theoretical linear elastic multilayer system model of the road a limiting design criteria of vertical compressive strains in the subgrade or in situ material is then used to assess the haul road under the specific loading conditions thereby determining the adequacy of the structural design vertical compressive strains induced in a road by heavy wheel loads decrease with increasing depth which permits the use of a gradation of materials and preparation techniques with stronger materials being used in the upper regions of the pavement the road as a whole must limit the strains in the subgrade in situ materials to an acceptable level and the upper layers must in a similar manner protect the layers below using this premise the road structure should theoretically provide adequate service over its design life figure 10 6 10 illustrates a mechanistically derived haul road structure in general terms applied load subgrade strength and the pavement structural thickness and layer strength factors predominantly control the structural performance of a haul road an upper limit of 2 000 microstrains is generally placed on subgrade or in situ layer strain values strain values exceeding 2 500 microstrains are associated with unacceptable structural performance in all but the most lightly traffic

ked and shortterm roads data from figure 10 6 2 can be used to assist in selecting a limiting strain value according to the category of road to be built and the associated operating life and traffic volumes in addition to prevent excessive damage to the wearing course deformation at the top of this layer must be limited to no more than 3 mm as for trolley assist roads haul roads in general have a much higher degree of traffic channeling in which trucks follow the same wheel paths so structurally no difference in the approach to structural design is warranted critical to the success of trolley assist is the height of the road surface under the trolley lines with a good structural design slight loss of running surface or compaction of the road structure will generally be within the limits of the truck pantograph system to prevent trucks being rejected from the trolley line grade changes under trolley should be avoided as far as possible however in the case of wearing course material management because of the high shear forces developed at the wheel road surface interface as trucks are accepted under trolley and accelerate severe wearing course raveling is often seen in these local areas this can be ameliorated to some extent by the use of well specified wearing course material and the use of dust palliatives which assist in binding the wearing course material mechanistic structural design procedure to determine the layer response to an applied load a layered elastic model can be used to represent the various haul road layers in the design software is available that can be used to solve multilayer problems in road design including elsym5a fhwa 1985 and circly mincad 2008 irrespective of the solution software used the approach is similar the effective elastic resilient modulus eeff and poisson s ratio typically 0 35 define the layerworks material properties required for computing the vertical strains v in a road in addition to the material properties a layer thickness 200 mm is also specified for the wearing course discussed later in the functional wearing course design section by varying the thickness of the waste rock layer a solution for maximum strain in any pavement layer that is below the limiting strain criteria for that class of road is found generally a three layer model is sufficient where the road is built directly on subgrade fill in pit blasted rock or in situ material ex pit softs or weathered overburden if the construction incorporates ripped and compacted in situ material this may also be added as another layer for computational purposes the layers are assumed to extend infinitely in the horizontal direction and the lowest pavement layer is assumed to be infinite in depth the applied load is calculated according to the mass of the vehicle and the rear dual wheel axle load distributions from which the maximum single wheel load is found the load application is deter

mined from dual wheel geometry and together with tire pressure the contact stress is calculated figure 10 6 11 summarizes the layered elastic model the strains induced in a pavement are a function of the effective elastic resilient modulus values assigned to each layer in the structure in order to facilitate a mechanistic design some indication of applicable modulus values is required figure 10 6 12 gives recommended modulus value correlations to the unified soil classification system uscs and american association of state highway and transportation officials aashto classification system to select suitable modulus values for in situ materials the associated range of cbr values alternatively derived in the field from dynamic cone penetrometer probing tayabji and lukanen 2000 are also given other values are published by thompson and visser 1996 and the south african roads board sarb 1994 equation 10 6 9 can also be used in conjunction with layer cbr values to determine modulus values eeff mpa but in each case care should be taken to ensure that the general correlations presented here are consistent with soil properties not directly assessed in the derivation of the equation the modulus value adopted for the selected blasted waste rock layer is typically 3 000 mpa this value is derived from consideration of a cement stabilized layer in its precracked large intact blocks with some shrinkage cracks state which corresponds closely to a well compacted waste rock layer method specification notes for structural design the structural design is based on the use of selected blasted waste rock this is subject to the material selected being blocky maximum block size 50 layer thickness and not containing excessive minus 25 mm sieve fines or clay contamination care should be taken when selecting blasted waste rock from weathered overburden will contain excessive softs to satisfy the layer specifications either the blast burden and spacing can be reduced or powder factor increased to produce a finer fragmentation than a conventional production blast alternatively a grizzly sizer can be used to remove the oversize material before placing in the road when constructing the various rock layers the blasted selected rock should be placed in a single lift if using impact roller compaction or 300 mm lifts if using vibratory roller compaction by running haul trucks over the road during construction end tipping and dozing to profile compaction of the layerworks will be improved and structural capacity increased in addition to improve compaction all excessively large boulders 50 of layer thickness should be removed if not scalped at the grizzly from the upper portion of the compacted rock layer the rock layer should be compacted dry by impact or vibratory roller compaction until no movement under the roller is visible this improves the structural response of the road to the applied loads and provides

an anvil for improved wearing course layer compaction following compaction a final pass of a grid roller can be used to assist in creating a correction layer of crushed or blasted overburden rock or similar can be placed if longitudinal or cross alignment needs correction fines should not be used for this purpose finally a gravel wearing course layer of 200 mm is placed onto the selected rock layer design of this layer is discussed in the following section functional wearing course design equally important as the structural strength of the design is the functional trafficability of the haul road this is dictated to a large degree through the selection application and maintenance of the wearing course or road surfacing materials poor functional performance is manifest as poor ride quality excessive dust increased tire wear and damage and an accompanying loss of productivity the result of these effects is seen as an increase in overall vehicle operating and maintenance costs the functional design of a haul road is the process of selecting the most appropriate wearing course material or mix of materials typically natural gravel or crushed stone and gravel mixtures that are commensurate with safety operational environmental and economic considerations the most common wearing course material for haul roads remains compacted gravel or gravel and crushed stone mixtures in addition to their low rolling resistance and high coefficient of adhesion their greatest advantage over other wearing course materials is that roadway surfaces can be constructed rapidly and at relatively low cost as with structural designs if local mine material can be used for construction the costs are all the more favorable this cost advantage is however not apparent in the long term if the characteristics of the wearing course material result in suboptimal functional performance an ideal wearing course for mine haul road construction should exhibit the ability to provide a safe and vehicle friendly ride without the need for excessive road maintenance adequate trafficability under wet and dry conditions the ability to shed water without excessive erosion resistance to the abrasive action of traffic freedom from excessive dust in dry weather freedom from excessive slipperiness in wet weather and low cost and ease of maintenance mcinnes 1982 thompson and visser 2000a the defects most commonly associated with mine haul roads in order of decreasing impact on hauling operational performance are typically as follows skid resistance wet skid resistance dry dustiness loose material corrugations stoniness loose potholes rutting stoniness fixed cracks slip longitudinal and crocodile crocodile cracks are cracks in a road caused by high plasticity high clay content material shrinking when drying this is a well known phrase among the pavement design fraternity method specification notes for functional desi

gn the method specification for placement of the wearing course depends on whether the haul road is being built from new or being rehabilitated where material is mixed with the existing wearing course to bring it back to specification for rehabilitation the existing wearing course layer should be ripped and scarified and any large lumps of compacted material broken down to a maximum size of one third of the compacted layer thickness during the processing the scarified layer should be ploughed or bladed to bring large lumps to the surface an offset disc harrow can be used for this purpose material to be mixed should be dumped opened and mixed with the existing wearing course placement of the wearing course should proceed in two lifts of 100 mm each compacted to 98 maximum dry density to give a minimum cbr of 80 using a large vibratory roller and with moisture content at or slightly dry of optimum material volumetric requirements must be determined such that the final thickness of the wearing course does not exceed 200 mm estimating haul road rolling resistance the rolling resistance of a haul road is primarily related to the wearing course material used its engineering properties and traffic speed and volume these dictate to a large degree the rate of increase in rolling resistance ideally road rolling resistance should not increase rapidly which implies that those road defects roughness defects leading to rolling resistance should also be minimized this can be achieved through careful selection of the wearing course material which will minimize but not totally eliminate rolling resistance increases over time or traffic volume haul road dust palliation dust generation is the process by which fine wearing course material becomes airborne such generation is termed a fugitive or open dust source the amount of dust that will be emitted is a function of two basic factors 1 the wind erodibility of the material involved and 2 the erosivity of the actions to which the material is subjected in broad terms the effectiveness of any dust suppression system is dependant on changing material wind erodibility or erosivity the wearing course silt and fine sand fractions i e between 2 and 75 m are a good indication of its erodibility the motivation for the use of some additional agent to reduce a material s inherent erodibility is based on increasing particle binding the finer fraction although contributing to cohesiveness also generates much of the dust particularly when the material is dry the presence of larger fractions in the material will help reduce erodibility of the finer fractions as will the presence of moisture but only at the interface between the surface and the mechanical eroding action this forms the basis of the water based dust suppression techniques used most commonly on mine haul roads the consequences of dust generation include loss and degradation of the road pavement material the fin

er particles being lost as dust and the coarser aggregates being swept from the surface or generating a dry skid resistance defect decreased safety and increased accident potential for road users because of reduced or obscured vision and reduced local air quality and higher vehicle operating costs vocs with dust penetrating the engine and other components resulting in increased rates of wear and more frequent maintenance many products available are claimed to reduce both dust and road maintenance requirements for mine roads often however no comprehensive comparable and controlled performance trials have been published additionally incorrect application techniques and construction methods often result which leads to considerable skepticism about such products and their overall cost effectiveness from a mining perspective the following parameters would define an acceptable dust palliative spray on application with deep penetration the ability to penetrate compacted materials or though less preferable mix in applications with minimal site preparation rip mix in and recompact straightforward applications requiring minimal supervision not sensitive nor requiring excessive maintenance or closely controlled reapplications the road trafficable within a maximum of 24 hours short product curing period availability in sufficient quantity at reasonable prices adequate proven or guaranteed durability efficiency and resistance to deterioration by leaching evaporation ultraviolet light and chemical reaction with wearing course or spillage on road effective over both wet and dry seasons evaluated against local and international safety standards and environmentally acceptable water based spraying a poor wearing course material cannot be improved to deliver an adequate performance solely through the addition of a dust palliative the haul road wearing course material should ideally meet the minimum specifications presented earlier if not the inherent functional deficiencies of the material will negate any benefit gained from using dust palliatives in road surfaces with too much gravel dust palliatives do not appear to work effectively more especially where a spray on technique is used as opposed to a mix in the palliatives do not aid compaction of the surface because of the poor size gradation nor form a new stable surface new surface area is created from exposed untreated material while with a mixin application poor compaction leads to damage and raveling of the wearing course traffic inducing breakdown of the material and eventual dust generation with regard to water soluble palliatives rapid leaching may be problematic in some climates in compact sandy soils tar and bituminous based emulsion products appear effective where leaching of watersoluble products may be problematic however in loose medium and fine sands bearing capacity will not be adequate for the tar bitumen products to maintain

a new surface and degeneration can rapidly occur in road surfaces with too much silt it is unlikely that a dust suppression program will be effective excessive silt or sand fractions may lead to a slippery road while poor bearing capacity leads to rutting and the need for road rehabilitation or maintenance which destroys most products small scale potholing has been observed on a number of pavements following spray on application or reapplication as a result of trafficking lifting fine cohesive material from the road again where no depth of treatment has built up this will lead to the creation of new untreated surfaces in general spray on applications do not appear appropriate for establishment of dust treatments especially with regard to depth of treatment required a spray on reapplication or rejuvenation may be more appropriate but only if penetration of the product into the road can be ensured otherwise it will only serve to treat loose material or spillage buildup which will rapidly break down and create new untreated surfaces a spray on treatment is however useful to suppress dust emissions from the untrafficked roadsides because it would be easier and less expensive to apply and with the material typically being uncompacted would provide some depth of penetration and a reduction in dust emissions because of truck induced air turbulence when chemical based dust suppressants are applied to an appropriate wearing course the average degree of dust palliation and the period over which it is applied is often seen to be considerably better than that achievable by water based spraying alone thompson and visser 2002 however in terms of cost effectiveness an evaluation is required with which to determine the extent of the cost benefits attributable to chemical based dust suppression together with an indication of those factors likely to alter the trade off between water and chemical based dust palliation a typical approach is illustrated in figure 10 6 14 finally it is worth noting that although tire chains are mostly limited to loading equipment when dust palliatives are used to maintain treated wearing course integrity all tracked and tire chained equipment should be kept off the road and either run along the road edge or be transported by lowbed trailers maintenance management design design and construction costs for the majority of haul roads represent only a small proportion of the total operating and road maintenance costs in particular the use of an appropriate road maintenance management strategy has the potential to generate significant cost savings particularly in the light of increases in rolling resistance because of the interactive effects of traffic volume and wearing course deterioration thompson and visser 2003 the management and scheduling of mine haul road maintenance has not been widely reported in the literature primarily because of the subjective and localized nature of operator ex

perience and required road functionality levels in most cases comment is restricted to the various functions comprising maintenance as opposed to the management of maintenance to minimize total costs long 1968 suggests that adequate serviceability functionality can be achieved by the use of one motor grader and water car for every 45 000 tkm metric ton kilometers of daily haulage the u s bureau of mines minerals health and safety technology division usbm 1981 in their report on mine haul road safety hazards confirm these specifications but without a clear statement as to what activities comprise road maintenance in addition to the lack of unanimous objectives in applying maintenance the definition of maintenance as applied to mine haul roads varies from mine to mine table 10 6 7 summarizes these maintenance activities modified for application to mine haul roads the routine maintenance category is adopted here to describe the various activities envisaged for haul road routine maintenance table 10 6 8 summarizes the various routine maintenance systems that mines typically apply routine maintenance is carried out on mine haul roads almost daily depending on the functionality of the road and the traffic volume the principal goals are to restore the road functionality to a level adequate for efficient vehicle travel with the aim of augmenting productivity and minimizing total road user costs and to conserve the integrity of the road wearing course by returning or redistributing the gravel surface ad hoc or scheduled blading is an inefficient means of road maintenance with the potential to generate excessive costs due to overmaintenance or undermaintenance of the road ideally an optimized approach is required with which to minimize total road user costs a maintenance management system mms for mine haul roads is described here following thompson and visser 2006 to address these needs maintenance management systems the ideal maintenance strategy for mine haul roads should be the one that results in the minimum total road user cost because in the case of mine haul roads the agency maintaining the haul road network is directly affected by road user operating costs two elements form the basis of road user costs namely road maintenance costs and vocs both these cost elements are directly related to road condition or more specifically rolling resistance the selection of a maintenance program for mine haul roads should be based on the optimization of these individual costs such that total vehicle operating and road maintenance costs are minimized as shown schematically in figure 10 6 15 mine haul road maintenance intervals are closely associated with traffic volumes with operators electing to forgo maintenance on some sections of a road network in favor of others this implies an implicit recognition of the need to optimize limited road maintenance resources to provide the greatest overall benefit this optimiz

ation approach is inherent in the structure of the mms for mine haul roads two elements form the basis of the economic evaluation namely 1 haul road functional performance rolling resistance increases with time traffic and 2 vehicle operating and road maintenance cost increase with time traffic the approach is suited to a network of mine haul road segments as opposed to a single road analysis for a number of road segments of differing functional and traffic volume characteristics together with user specified road maintenance and voc unit costs the model assesses traffic volumes over network segments during the analysis period as specified the change in road as rolling resistance by modeling or on site assessment the maintenance quantities as required by the particular strategy the vocs by modeling total costs and quantities and the optimal maintenance frequency for specified segments of the network such that total road user costs are minimized figure 10 6 16 illustrates the mms flow chart used to determine optimum maintenance interval for a mine road consisting of a number of road segments cost savings associated with the adoption of an mms approach are dependant on the particular hauling operation vehicle types road geometry tonnages hauled and so forth the first element of an mms for mine haul roads is based on modeling the variation of vocs with rolling resistance when combined with a road maintenance cost model the optimal maintenance strategy for a specific mine haul road commensurate with lowest overall vehicle and road maintenance costs may be identified vehicle operating cost model the voc model refers to the incremental cost of truck operation with changes in road rolling resistance the cost model should consider the effect of increased rolling resistance on fuel consumption tires and vehicle maintenance however a reasonable approximation can be determined from fuel consumption alone the prediction of fuel consumption variation with rolling resistance involves simulation with specific haul trucks to generate a speed model for various road grades the speed model forms the basis of the fuel consumption model derived from vehicle simulations coupled with vehicle torque fuel consumption maps road maintenance cost model the road maintenance operating cost per kilometer comprises both grader and water car operating costs although not contributing directly to a reduction in rolling resistance the incorporation of the watering costs in the maintenance costs model reflects the ideal operating practice in which immediately before blading the section of road is watered to reduce dust and erosion and to aid blading and recompaction grader and water car productivities of 0 75 and 6 25 km maintained road per operating hour for each machine respectively are typical and correlate with published figures of between 8 and 18 km of maintained road per 16 hour day however as the condition of the h

aul road deteriorates maintenance becomes more time consuming and the number of blade passes required to achieve an acceptable finish when the rds exceeds 45 equivalent to approximately 3 rolling resistance increases a maintenance productivity curve shown in figure 10 6 17 incorporates this reduction in grader productivity associated with excessively rough roads rds can be converted to rolling resistance through equations 10 6 16 to 10 6 18 the road maintenance cost model is thus constructed from consideration of the average blade width per pass road width rds before blading motor grader productivity curve and hourly cost from which the motor grader cost per kilometer is found this cost is then combined with the cost per kilometer of the water car and workshop costs to produce a total cost per kilometer for road maintenance estimating rolling resistance from road visual assessments in addition to the estimation techniques for rolling resistance progression presented earlier in the chapter rolling resistance can also be estimated from qualitative visual evaluation a road defect classification system can be applied in which the key defects influencing rolling resistance are identified and the product of defect degree measured from 1 5 and extent measured from 1 5 are scored following table 10 6 9 the sum of the individual defect scores thus rated equivalent to the rds discussed previously can be converted using figures 10 6 18 and 10 6 19 to give a rolling resistance for the segment of haul road under consideration more details of the approach and aids to visual evaluation of defects are given by thompson and visser 2006 after the mineral products are extracted from an underground or surface operation they need to be moved stored and recovered before being treated further or sold to external clients traditional and small operations are based on small scale discontinuous movement of materials that are labor intensive the trend of the industry is toward larger continuous systems with increased automation levels while in the first part of the 20th century a 1 000 t h metric tons per hour 1 100 stph short tons per hour operation could have been classified as large volumes of 10 000 t h 11 023 stph are quite common today some continuous operations reach 40 000 t h 45 000 stph and several mining operations reach levels close to 1 mt d million metric tons per day 1 1 million stpd between ore and overburden removal in most open pit mines the volumes of overburden and waste rock required to be moved are much larger than the ore volume ratios between waste and ore of 2 to 6 are quite common they can get as high as 30 before an underground operation becomes a better option the old paradigm of abundant and cheap labor in third world countries is disappearing globalization brings common standards for environmental and safety requirements at the same time that economic development reduces the income gap b