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y designation rqd values at the least and any rock mass classification work that has been done results of any special studies or examinations the exploration department has performed metallurgical tests geotechnical work etc report on any special problems or confrontations with the local populace any other pertinent data such as attitude of local populace toward mining special environmental problems availability of water and hydrologic conditions in general and infrastructure requirements ideally a number of mining and processing alternatives will be examined as a screening process obviously these need not be in depth studies but most experienced mining engineers will quickly be able to determine what mining methods will be applicable and can then place costs on several alternatives for this application likewise an experienced mineral processor can determine the candidate process flow sheets and can place costs on these alternatives at the same time all the other elements of the project must be considered and studied in just enough detail to discover any fatal flaws or problems that need engineering mitigation certainly environmental and socioeconomic issues need to be studied and scoped to the extent that any existing or expected problems will be detected then all of these items can be examined for future cost and work plans costs and expenditure schedules will be based on industryfactored historical experience major capital costs can be based on telephone quotes from suppliers or canned commercial programs built for this type of application usually no field work or metallurgical testing will be conducted unless a definite metallurgical problem is recognized with the resource and suspected to be a fatal flaw in which case it should be studied depending on the complexities of the project approximately 5 000 to 30 000 worker hours of work is needed to complete these activities during the preliminary study this description is written for a company or group that is prepared to perform most of the evaluation activities with various contractors thus for every task that is to be contracted 1 a scope of work must be developed 2 the industry needs to be surveyed for potential contractors 3 contractors must be evaluated to ensure they are qualified and likely to perform as expected 4 a request for bid proposal documents needs to be prepared and sent out 5 bid documents must be evaluated and the award made 6 negotiations with the winning bidder may be necessary if there have been variances to its bid package and adjustment made and finally 7 the contractor can be mobilized all of these take a considerable amount of effort if the work is to be done by experienced in house engineers on the project team or from that function of an organization then the contracting procedure does not apply however such activities as writing the scope of work should still be carried out by the central project team to make

sure that potential challenges problems are fully identified and that potential impacts are considered for all the other parts of the project the results of this preliminary study will be adequate for comparative screening of mining and processing alternatives while an economic analysis will determine whether to proceed with or reject the project a primary objective of the study is to plan and estimate costs for a further predevelopment program if warranted approximately 4 to 8 of the project engineering will need to have been completed in which case the probable error of cost estimates accuracy should be between 35 and 45 if 10 to 12 of the total project hours have been completed in the study the probable error should be between 30 and 35 while contingencies of between 20 and 35 for capital costs will apply an economic analysis will be performed and the preliminary feasibility report will be fully documented at this point presentations to management will be made and depending on the results of the economic analysis approval to proceed to the next step of the project or otherwise follows major activities of the preliminary feasibility study a description of these preliminary study activities and tasks can be found in appendix 4 7a this generic list applies to all mineral properties but can be adapted with addition and deletions for a particular deposit remember for any of these activities that need to be contracted the seven contracting steps listed will consume a lot of time phase ii intermediate feasibility study based on results of the preliminary study showing that a project has the potential to achieve the desired company goals the intermediate feasibility study should be initiated the specific object for this study is significantly different than for the preliminary study now that it has been shown by using at least one mining processing system that the mineral resource being examined has potential economic viability the objective must now focus on methods to optimize each component mine plant process while at the same time taking an in depth look at all of the project parameters briefly studied in the preliminary study at this time accurate topography maps specific to the area must be generated if not already available any shortcomings in the land and water status discovered in the preliminary study must be corrected at this point before investing any more money mine design will be based on information from the early exploration delineation drilling program plus any additional exploration sampling done between the two phases in some cases bulk sampling may be required thus if permits can be obtained a test mine may be justified after this phase of the study if further exploration drilling or trenching takes place during this phase permits and contractor agreements must be prepared under the control of the project team the sampling program must prepare a sample flow chart prep

are a chain of custody security procedure if not already in place designed to protect the integrity of the eventual sample analysis and procure and analyze the new samples the new geology and mineral information must be fed into the database and evaluated after rebuilding and analyzing the new database and documenting the current reserves and resources new reserve and resource maps can be constructed for mine planning given the shape and character of the ore reserve identified to this point the mine planning will begin only measured and indicated geologic resource material may be used for mine planning in the united states and canada although those mining methods considered in the preliminary study may be reexamined other methods should also be considered since the ore body shape size character and grade may have changed the methods described in this text on mine planning and mining methods should be followed but this time after a rough screening of multiple mining methods two or three of the most probable mining methods or variations which are considered safe and environmentally permissible and that will probably yield the lowest cost or greatest recovery should be carried through the study until an economic comparison can be made likewise with the latest mineralogical data and mining methods several mineral processing and waste disposal alternatives should be considered and those that seem likely to yield the best economics should be carried through the study until a true economic comparison can be made between the methods facilities siting and geotechnical investigations will need to be conducted if competent personnel are not available within the company contract preparation to cover the scope of work for the approximately 125 intermediate feasibility activities must be done the same list of contracting activities must be completed that are shown in appendix 4 7a for phase i the preliminary study and time must be scheduled for all of this contracting effort one must not underestimate the time it can take to perform these tasks scopes of work requests for proposals rfps survey of industry contractors obtaining a legal contract allowing time for bids and evaluating the bids and negotiating the contract details with the successful contractor this must be done for each contractor it may take several months to get a contractor assessed and the details of their contract accepted but with contractors on board work can then begin with environmental baseline studies impact assessments and long lead time permit applications again if competent personnel are not available within the company contract preparation to cover the scope of work for these environmental activities must be done bidding lists must be prepared rfps issued bid evaluation criteria written and the bid evaluation administered finally when negotiated details of the contract are completed environmental contracts are awarded a

lthough the baseline studies take time they should be completed by the time the intermediate feasibility study is done which will allow for this information to be submitted along with the intermediate mining and process planning to the permitting agencies results of the intermediate study will be adequate for determining economic feasibility and defining additional predevelopment and or metallurgical testing requirements in many cases the benefits and requirements for a test mine or bulk sampling will be fully recognized and defined at this point in most cases specific permitting will be required and this will require time to receive such permits the cost estimates for the two or three alternatives developed during this phase should be based on detailed functional analysis of the mining and processing methods of each operation on suppliers written quotes and on bench scale metallurgical testing by the end of the intermediate study the engineering on the project should be between 12 and 15 completed the probable error of cost estimates should be 15 to 20 while contingencies of between 15 and 20 will apply economic analysis will be performed on the favorable sets of alternatives selected usually no more than three sets of alternatives will be evaluated when the intermediate feasibility report is fully documented presentations to management will be made and depending on the results of the economic analysis approval to proceed to the next step of the project or otherwise will follow major activities of intermediate feasibility study a description of these intermediate study activities and tasks can be found in appendix 4 7b this generic list applies to all mineral properties but can be adapted with addition and deletions for a particular deposit for any of these activities that need to be contracted the seven contracting steps listed in the preliminary feasibility discussion apply and will consume a lot of time need for a test mine many times the feasibility team will not be able to obtain enough ore quality and geotechnical information working with just exploration sampling in these cases a test mine must be considered the test mine may come after the intermediate feasibility study or during the final feasibility study the advantages of a test mine are as follows from a mining perspective verifies the expected ore continuity thus eliminating disastrous surprises accurately assesses the rock strength allowing prudent planning and sizing of the commercial mine opening verifies mining efficiency and productivity as it relates to drilling blasting and materials handling determines from reliable water studies the nature of mine water inflows thus allowing for adequate water handling procedures to be installed before problems are encountered better quantifies the mine ventilation friction factors and requirements and confirms the character of the waste product and how it will be handled in the comme

rcial operation from a metallurgical perspective verifies and optimizes the metallurgical flow sheet with a pilot plant process that is continuous lock cycle testing determines what size and type of equipment will be optimal for the metallurgical recovery determines what type and amount of reagents will lead to the best recoveries and concentrate grades determines the required amount of water and how to achieve a water balance provides a more accurate prediction of concentrate grade moisture content and impurities and provides a much better assessment of the work index from a bulk sample than from small samples from an environmental perspective demonstrates the ability to control the operation in such a manner that it will not harm the environment allows the project team to completely study the waste characterization and determine any future problems and if water discharge is involved allows the project team to study the difficulty of settling the discharged water and determine what is necessary to mitigate future problems and determine if zero discharge is possible from an engineering design perspective improves the ability to make more accurate cost estimates because of better knowledge of the abrasivity of the rock and of the ground slope control of the stopes pit walls which could lower the cost estimate because less contingency may be needed improves labor estimates because of a better understanding of the productivity of each unit operation predicts a more accurate schedule because of better understanding of the unit productivities and lowers the overall risk of the project in every aspect from the perspective of expediting later mine development shortens the overall schedule from the end of the feasibility study to the end of construction because of early access to develop the commercial mine and completely utilizes the openings as part of the commercial mine operation because access to the underground opening already exists some shafts may be able to be raise bored and then expanded by mechanical excavation rather than by the more expensive conventional shaft sinking methods and finally the test mine may be an ideal training facility prior to the commercial mine start up activities related to the test mine are found in appendix 4 7c phase iii final feasibility study the final feasibility study should be initiated when results from the intermediate study show that the project still has the potential to achieve the desired company goals the objective as in the first two phases is still to determine the potential value of the property to its owners either by determining the optimum method of developing it by selling it or by doing nothing further at the moment however more specific to the final study the objective now becomes one of minor refinements to all of the details of the intermediate study that yielded results that met the company objectives this is designed to optimize the return on

the future investment the final feasibility study will be prepared during the permitting time schedule for the project since final project features must be reflected in the permits to accurately assess impacts and mitigation for the agencies commentators stakeholders nongovernmental organizations ngos and community assuming that this project still shows favorable results at the end of this phase of study the design parameters set in the final feasibility study will feed into the design basis report which guides the project into the design and construction phase and finally into operations if test mining with bulk sampling and pilot plant testing has not been completed it now becomes part of the final feasibility study mine and process facilities will be further studied and the best alternative developed in the intermediate study will be optimized using the latest exploration and metallurgical test data probably from the test mine bulk sample the reserves will be updated and the metallurgical flow sheet will be optimized final environmental impacts will be determined following prescribed guidelines applications for construction and operating permits will usually be made early in this phase of study subject to later modification mine and process operating cost estimates will again be made by performing a functional analysis capital cost will be refined by again soliciting written quotes from vendors by the end of the final study the engineering should be 18 to 25 completed the probable error of cost estimates should be 10 to 15 of the total cost estimate with a contingency of 10 for most engineered structures other less well defined aspects of the project e g mine development should have contingencies of at least 15 an economic analysis will be performed with the final feasibility report fully documented at this stage presentations are made to management and depending on the results of the economic analysis it will give approval to proceed to the design and construction phase of project development and for the associated budget major activities of final feasibility study plan and budget a description of these final feasibility activities and tasks can be found in appendix 4 7d this generic list applies to all mineral properties but can be adapted with addition and deletions for a particular deposit for any of these activities that need to be contracted the seven contracting steps listed in the preliminary feasibility discussion apply and will consume a lot of time combining classic approach with recommended approach because there are good reasons as outlined at the beginning of this chapter to sometimes utilize the nonuniform classic conceptual scoping study approach if the decision is made to move the project to the next level of study one should convert to the engineered systematic three phased approach in this case the conceptual scoping study should be compared with the details of the

engineered systematic phased preliminary study whereas in the classic scoping study only 2 of the engineering may have been done for the first phase in the engineered systematic preliminary study as much as 4 to 8 of the engineering would be completed this means that if one is now going into the intermediate feasibility study in the engineered systematic three phased approach the extra work to be done in the intermediate study must be planned so that by the end of the intermediate study all of the reasonable functional alternatives of the operation should have been examined the optimal method selected and 12 to 15 of the engineering work completed before going into the final feasibility study most mine management people believe that the final feasibility study is the final document of the feasibility process unfortunately with this in hand most management teams will head for the bank or a joint venture partner and on to a design contractor without fully documenting everything that went into the full feasibility study this includes not only a description of the geologic reserves mine and plant facilities and hardware along with the positive economic picture but all of the company design and operating philosophy that is so important when the build and operate stage is reached the plans on how to execute the project and how to operate the completed facility are equally as important to the design contractor the banker and the potential joint venture partner if all important activities through to the final operation have not been examined and then documented the project is not ready to move forward project design basis report preparation with project approval the course of action will either be to go directly into the project execution phase or possibly seek out a joint venture partner in either case the design basis report dbr sometimes called a design basis memorandum or design basis document needs to be prepared the approved project feasibility report will be presented in sufficient detail to produce a dbr which is the document that will guide the project through the next step designing the project based on the preceding studies why a dbr is needed the primary purpose of the dbr is to be able to convey to future design engineers a consolidated document in which all the needed information is contained in a condensed version but it can also serve to inform others such as financial organizations construction personnel or persons who may be interested in a joint venture although much of the information is also contained in the final feasibility study this document is written more for the purpose of documenting for management that the project is indeed both feasible and economically viable in contrast the dbr is written to convey all of the technical information that will be needed by the architect engineering a e design organization which has already been developed by the owner s project feasibilit

y team it contains all of the drawings prepared during the final feasibility study plus any others required to convey the needed technical information to the a e organization for the project and will be used as a basis for the final bids by the various a e organizations in the introduction the writer should define the purpose and use of the dbr at a minimum the dbr serves several purposes the dbr defines the technical basis for project design and construction so that basic detailed engineering downstream can proceed provides the basis for a coordinated review by the organizational entities involved the future operations group the engineering group management and the future a e team provides documentation for the technical basis and facilities description from the final feasibility study cost estimates and conveys the company s construction and procurement philosophy to the future a e team the dbr is usually written in several volumes in the case of the following example five volumes were produced rather than describing in specific detail what should be written under each section and subsection of each volume a brief description is given here concerning the general content of that volume with a generic outline of items given in appendix 4 7e volume 1 management summary the management summary prepared by the project executive or project manager as applicable summarizes the project objectives the assumptions that were made the work that has been completed the economic analysis and associated risk and the recommendations of the project team other items that should be covered if studied by the project team include project funding and the business plan with market and competition analysis and strategies any outstanding major issue involving government agencies related to utilities transportation land royalties or potential project partnerships must also be mitigated at the end of the summary are the conclusions and recommendations with discussions on the reserves the feasibility of the project the market the schedule of the design construction and start up as planned in the feasibility study any preappropriation work contemplated and needed funding an example outline of the information that is contained in volume 1 management summary is shown in appendix 4 7e volume 2 project economics prepared by the project executive or project manager project economics summarizes the capital and operating costs project schedule market forecasts inflation projections if constant dollar analysis was not used and other factors that affect the total erected cost and project economics identified project risks and the measures needed to mitigate those risks should be documented an example outline of the information that is contained in volume 2 project economics is shown in appendix 4 7e volume 3 technical narrative prepared by the project team the technical narrative describes the technical basis fo

r the project and lists the design considerations and constraints this is the technical meat of the project the narrative must convey to the future a e contractor exactly what is to be built and exactly what the a e constructor is to do and precisely how it will accomplish that nothing can be left out for this reason all of the drawings prepared during the final feasibility stage plus whatever drawings are necessary to convey the message to the a e organization must be in the dbr the better defined the project is in the dbr the more accurate the cost will be to the bid estimates and the fewer exceptions that will have to be negotiated an example outline of volume 3 technical narrative is found in appendix 4 7e volume 4 project execution plan this document prepared by the project team defines the real and potential problems in the detailed engineering procurement and construction of the project furthermore it goes on to describe the best plans to ensure that these problems are mitigated or at least minimized the recommended contracting plans are spelled out as are the plans for engineering and design procurement and construction an example content outline of volume 4 project execution plan is in appendix 4 7e volume 5 operating plan prepared by the company s operations department the operating plan explains how to minimize the impact for identified potential problems in start up and continuing operations although much of the company s operating philosophy should already have been placed into the design as presented in the final feasibility study the writers of the operating plan should again emphasize the company s attitude toward mechanization and automation and what they are willing to pay for it and its policy on safety and environmental issues and maintenance and contracting such issues as labor recruitment and training will be planned scheduled and budgeted learning curve estimates will be applied toward the production buildup so the estimated production will be on schedule and project economics will be preserved an example outline of volume 5 operating plan is found in appendix 4 7e it is this dbr document that is used as the basis for the subsequent engineering design not only does it contain the technical data and information decided on by the company during the final feasibility study but also the project execution plan for contracting building and constructing the mines of the project it also contains the operating plan which will guide the engineers and builder to construct the mine plant so the operating philosophy of the company can be quickly achieved and maintained historic information on the activity duration that goes into the project schedule and the functions that will be performed in the engineering design and final constructed project is included in the next section it often appears to laypeople that building mines and plants takes much too long and costs too much but this is

not the case what is true is that the expectations based on most of the final feasibility studies are overoptimistic and thus the project begins badly data from case histories are presented in the next section showing what should be expected feasibility and project timing and schedule the time it takes between the discovery of a resource that may be a potential ore body and when the ore body is brought into production can vary significantly obviously with an extremely high grade ore body it may take significantly less time to identify enough ore to start mining likewise if money can be made no matter how you mine it then the company may not want to spend a lot of time optimizing the mining and milling methods though this could and has proved to be a mistake in the past on the other hand it may take many years to define large marginal mineral resources and to optimize every aspect of the study in order to turn the resource into a viable reserve technological changes over time may also allow the property to be developed after many years of study the other factor is the ever changing environmental permitting which can vary significantly although a small punch coal mine in appalachia can take less than a year the development of a world class zinc copper mineral resource in wisconsin united states was stopped for 20 plus years and finally terminated even though it can be demonstrated that the underground mine can be built and operated in a manner that would be environmentally acceptable anywhere else overall it usually takes from 2 to 6 years just to complete the mineral property feasibility evaluation study the overall time from find to mine is logically divided between the classical phases of mineral development preliminary exploration and discovery land and water acquisition commitment exploration feasibility studies environmental permitting final engineering development and construction start up to full production all of these activities vary greatly in length in this chapter only the length of the feasibility studies and environmental permitting will be considered nelson associates nelson 1979 conducted a study for the u s bureau of mines on the duration of these phases of mineral development for four mining projects in wisconsin and minnesota united states three were operating and one was being considered because these states have strict systems of environmental permitting for new mine development it is not surprising that the nelson study found that permitting time for a metal mine in wisconsin was by hindsight more than 100 optimistic for example the following information comes from nelson s summary of time table environmental monitoring 2 18 years environmental impact report evaluation 4 85 years state permits 5 25 years local permits 3 25 years environmental impact report preparation 1 55 years wisconsin department of natural resources wdnr for the wisconsin en

vironmental protection agency 3 90 years federal environmental impact statement eis 3 05 years master hearings 1 16 years other wdnr permits 3 05 years in reality some of these activities can go on simultaneously or overlap but even the most optimistic schedule to receive permits in wisconsin for a metal mine was 10 to 12 years in addition in the same report nelson also developed a time estimate based on these four major projects for feasibility studies as 8 years which seems too long however this information came from four major projects done by four different major mining companies although it certainly can take 8 years to do a complete mineral property feasibility study for many projects it is not necessarily true for all of them table 4 7 3 is taken from one mineral company s estimates of the average time expected to complete the feasibility evaluation on 10 small to large projects however this schedule overlaps all activities possible on maintaining a high level of engineering standards the approximate 300 total activities listed previously completed mostly by contractors depending on the size grade location ownership of the project and how much financing the owner needs these times can change radically the total worker hours to complete a feasibility study from the conceptual study through the final feasibility study will vary from 50 000 worker hours to more than 100 000 worker hours by the time the final feasibility study is finished 18 to 25 of the total engineering for the project will have been completed in contrast to these long time estimates cusworth 1993 presented the estimates for australia as typical durations for feasibility studies scoping study 7 to 9 months prefeasibility study 9 to 13 months feasibility study 12 to 17 months therefore it could be concluded from cusworth that all projects in australia vary only from a total of 28 to 39 months unfortunately no details are given as to what was actually covered during these periods it would have to be assumed that much of the difference between the united states and australia is the u s environmental agencies red tape but two other factors may play a significant role 1 probably more virgin deposits were being discovered in 1993 in australia than in the united states which might tend to be of a higher grade and 2 australians tend to turn everything over to contractors which with their larger staffs can usually perform faster than their u s counterparts scheduling of each project element must be done from the beginning this is one of the important reasons to document in advance all of the activities of each feasibility level phase then estimated worker hours of labor time must be assigned to each of these hundreds of project activities and subactivities setting up and maintaining the schedule of even a medium sized project is a major task involving thousands of activities organizing the project team

there are many ways to organize a project team depending on phase of the feasibility study size and complexity of the project location of the project and size and experience levels within the parent company first the talent that is needed either part time or full time on a project feasibility team must be considered certainly the team needs people who understand and can perform project management as well as costing and scheduling for the project in addition every technical discipline that has been considered in the evaluation must be populated this includes the fields of geology geostatistics mining metallurgy environmental consideration hydrology geomechanics civil infrastructure and economic evaluation but there must also be people who can provide and evaluate legal land water public relations socioeconomic marketing tax and financial information depending on the size of the parent company it must either build the organization within the company structure or depend on the consulting industry to supply the needed talent taking the in house approach the company must form a project management and development organization whose staff will be assigned to the project management nucleus of each project in addition technical specialists are assigned from a technical support organization on an as needed basis to perform the multitude of technical activities that will be required by approaching the problem in this fashion and using proper labor scheduling many projects can be handled simultaneously this approach works well on small to medium sized projects up through the intermediate feasibility phase of study for large or mega projects it would probably only work in the preliminary phase taking the consultant approach the company should still form a project management organization to manage each project but then contract to either one large multidisciplined consultant organization or individual discipline consultants to perform the various technical tasks of each project the consultant approach will not be discussed in detail since the a e consultant normally supplies all the organization s needs in house project teams two basic management philosophies must be considered functional matrix or line staff organizations this operating philosophy goes beyond the consideration of managing projects but is the means by which all supplied labor report to the company s various functional departments there will be a considerable difference in how the project teams are made up depending on whether the company operates as a line staff organization a functional matrix organization or a hybrid of the two it is difficult to generalize but if the company is running several small to medium sized projects which are in the preliminary or even the intermediate feasibility study phase it is more cost effective to organize a core group consisting of the project manager and a project cost and schedules coordinato

r and temporarily assign experts from the matrix technical organization rather than organize several line staff organizations for each project by allocating work in this manner each discipline can usually handle several projects at one time with proper scheduling it can usually work well through the intermediate level of feasibility particularly if the projects are located in the same country as the home office mine evaluation and development group however when the project is overseas or if it is a large project the amount of field work required during the intermediate and final studies mandates that it is usually best to move the dedicated project team to a location near the site at such time as the magnitude and importance of a project justifies it a separate project team organization is established this would normally occur at the end of either the preliminary or intermediate study phase when the cost duration level of staffing or overall importance indicates that a separate project team should be created at this point the team would be transferred to a location close to the site of that project but still functioning under the project development group it is also worth noting that in some countries there is little or no technical mineral engineering base on which to draw for a staff and one must be literally imported at the conclusion of the preliminary or intermediate study phase and when it is determined that a separate project organization should be established an independent project team is organized and works through the local organization as determined by management and coordination with the affiliate or country manager in such cases the project may have a project executive who also serves on the local country management committee or staff in house project teams two basic management philosophies must be considered functional matrix or line staff organizations this operating philosophy goes beyond the consideration of managing projects but is the means by which all supplied labor report to the company s various functional departments there will be a considerable difference in how the project teams are made up depending on whether the company operates as a line staff organization a functional matrix organization or a hybrid of the two it is difficult to generalize but if the company is running several small to medium sized projects which are in the preliminary or even the intermediate feasibility study phase it is more cost effective to organize a core group consisting of the project manager and a project cost and schedules coordinator and temporarily assign experts from the matrix technical organization rather than organize several line staff organizations for each project by allocating work in this manner each discipline can usually handle several projects at one time with proper scheduling it can usually work well through the intermediate level of feasibility particularly if the projects are located

in the same country as the home office mine evaluation and development group however when the project is overseas or if it is a large project the amount of field work required during the intermediate and final studies mandates that it is usually best to move the dedicated project team to a location near the site at such time as the magnitude and importance of a project justifies it a separate project team organization is established this would normally occur at the end of either the preliminary or intermediate study phase when the cost duration level of staffing or overall importance indicates that a separate project team should be created at this point the team would be transferred to a location close to the site of that project but still functioning under the project development group it is also worth noting that in some countries there is little or no technical mineral engineering base on which to draw for a staff and one must be literally imported at the conclusion of the preliminary or intermediate study phase and when it is determined that a separate project organization should be established an independent project team is organized and works through the local organization as determined by management and coordination with the affiliate or country manager in such cases the project may have a project executive who also serves on the local country management committee or staff 3 apply equipment costs wages salaries and supply prices to the cost parameters to estimate associated mine capital and operating costs 4 compare estimated costs to the anticipated revenues under economic conditions pertinent to the project using discounted cash flow techniques to determine project viability after the estimator evaluates the results he or she will make adjustments to the design and the production rate as necessary and then repeat the process preliminary mine design the goal of the mine planner is to optimize economic returns from the deposit or to otherwise achieve the corporate goals of the project s owners the objectives of evaluators as they design a mine for the purpose of estimating costs is to determine the equipment labor and supply requirements both for preproduction development work and for daily operations the extent to which the evaluator takes the design is important the process is one of diminishing returns roughly speaking 10 of the engineering required for a complete mine design probably provides the data necessary to estimate 90 of the costs more detailed final engineering aspects of mine design such as those needed to ensure adequate structural protection for the workers and sufficient ventilation of the underground workings seldom have more than a minor impact on the overall mine costs at the initial stages of an estimate the key element is distance in the preliminary design engineers need to establish the critical distances associated with access to the deposit whether by shaft adit or ra

mp most of the costs associated with preproduction development are directly tied to the excavations required to access the deposit the length or depth of these excavations along with their placement provide several cost parameters directly such as those needed to determine preproduction consumption of pipe wire rail and ventilation tubing these distances also provide an indirect path to estimating preproduction consumption values for items such as explosives drill bits rock bolts shotcrete and timber and finally they impact many subsequent calculations that the evaluator must undertake to estimate the required sizes of pumps ore haulers hoists and ventilation fans engineers let the configuration of the deposit and the structural nature of the ore footwall and hanging wall dictate the stoping method used to recover the resource stoping method selection is discussed in great detail in other chapters of this book the underground development openings necessary to access and support the stopes are as important as the stoping method itself engineers rely on some basic calculations to estimate the lengths of the drifts crosscuts ramps and raises associated with each stope after they determine the amount of ore available in the stope they use those lengths to approximate the daily advance rates needed to maintain the desired ore production rate while it is true that the use of average rates can be quite misleading particularly in the first 5 years of operation for the purposes of estimating costs in particular in estimating preliminary costs the overall costs per ton will not change much through the process scheduling of these activities in detail and while the timing of the costs will have some impact on project economics the extent to which they are detailed vs the overall impact on project economics refers back to the statement about diminishing returns in the first paragraph of this section in the process of determining stope development requirements estimators rely on stope models in conjunction with the deposit dimensions for example the diagram in figure 4 8a 1 appendix 4 8a provides the basis of the stope design for the room and pillar method from that basis the evaluator can then move on to use relationships similar to example 1 to establish the design parameters for the stope sketches usually a three view drawing of the deposit access headings stopes and underground excavations shops pump stations lunch stations hoist rooms etc provide much of the preliminary mine design information that an evaluator needs for a cost estimate only the lengths of the excavations are needed early in the analysis values determined by evaluators as they calculate the subsequent cost parameters provide the information necessary to define the cross sectional areas of these openings cost parameters engineers find that the process of defining the parameters necessary for a cost estimate is a wonderful perh

aps only for an engineer progression of simple mathematical calculations in which one value seems to always lead to and interconnect with the next these calculations branch in ways that create many logical paths to a complete compilation of the needed cost estimation parameters but all paths do eventually lead to such a compilation one generalized path is illustrated in this section be aware however that most of these procedures are interchangeable and many paths exist also it is not the intent here to work through a complete step by step estimate because such an example would apply only to a finite number of deposit types the intent is instead to provide insight into the process that estimators use and to remove some of the mystery that might create a hesitation to proceed successful estimators need to show a willingness to suggest values for as of yet unknown parameters for example if one is working at an operating mine then it is possible to know all of the required parameters and to calculate as opposed to estimate the costs but for undeveloped projects it is not possible to know parameters such as the ore and waste powder factors and the amount of water that must be pumped on a daily or hourly basis most if not all parameters required for a cost estimate fall into one of three categories in that they define labor supply or equipment requirements these categories represent the items that cost money that is the items for which funds must be expended consequently evaluators work in this phase of the estimate to specify the equipment the supplies and the work force necessary to mine the deposit they find that the key to specifying these factors lies in the process of determining how much time how many hours it takes to perform the individual tasks of mining operations in an underground mine are for the most part either cyclic or continuous in nature and most are designed to transport materials such as ore waste air water workers and supplies operations that do not transport materials such as equipment repair or rock support installation are typically in place solely to service operations that do the rate at which ore is produced provides evaluators a good place to start as they begin to define the cost parameter values this rate is typically based on the desired life of the mine and the size of the resource resource size is known or it has at least been approximated hence the evaluation and estimators often begin with the following relationship a variation of which is known as taylor s rule to approximate a possible project life of course many factors influence the rate of ore production such as market conditions deposit configuration and profit maximization so many evaluators use more sophisticated approaches to determine the initial production rate tatman 2001 as the evaluation of a project proceeds however the production rate needs to be altered from one iteration to the nex

t as the economic ramifications of each development scenario become clear after an initial production rate is determined evaluators can use it in conjunction with the ore haul distances gleaned from the mine design to estimate the capacities of the machines used to collect ore in the stopes transport it through crosscuts and drifts and then finally haul it through an adit ramp or shaft to the surface the heights and widths of these machines rear dump haulers scoop trams rail cars conveyors etc provide the basis for the cross sections of all the openings through which they must travel evaluators multiply the products of the heights widths and lengths of the openings by the density of the rock through which they pass to determine the amount in metric tons of rock that must be removed during their excavation when they apply a powder factor kilograms of explosive per metric ton blasted to this amount the result is the amount of explosives needed to liberate the rock as stated earlier evaluators find that as one design parameter is determined its value usually provides the information needed to determine many more after they determine the size of the haulers engineers can refer to manufacturer s literature often available through their web sites to ascertain the speeds of those machines in relation to various haul conditions and with those speeds they can calculate how many hours the machines need to operate each day to meet production goals and in turn the required number of machines and operators evaluators rely on cycle time calculations example 2 to supply the basis for most such values and as such these calculations represent one of the more important concepts of any cost estimate cycle time calculations are used whenever an estimator needs to determine the number of machines required to perform a cyclic operation example 2 cycle time calculations consider a case where a 20 t capacity articulated rear dump truck hauls ore to the surface ore is placed in the truck by a 6 1 m3 capacity remotely operated loader near the entrance of the stope the truck hauls the ore 550 m along a nearly level drift and then it hauls the ore 1 450 m up a 10 gradient to the surface after reaching the surface the truck travels another 200 m to the mill where the ore is dumped into a crusher feed bin if the truck is not loaded to capacity it is primarily because the capacity of the loader bucket in conjunction with the number of cycle either under or overloads the truck cost estimates after an evaluator has established all of the cost parameters the estimation process is one of simple calculations and tabulations because most costs both capital and operating are tied to average daily equipment use supply consumption wages or salaries estimators from this point forward need only identify the most reliable source of cost information apply the costs to the previously derived parameters and then tabulate th

e results this process is well demonstrated in the cost models found in appendices 4 8a through 4 8c daily equipment use varies greatly according to the mine development schedule so the estimate must be based on snapshots of the schedule at representative times that is this snapshot is the goal evaluators calculate equipment operating costs by multiplying use in terms of hours per day by the hourly operating costs for the machine which they typically glean from sources that include the mine and mill equipment operating cost estimator s guide infomine usa 2009a and the cost reference guide equipmentwatch 2009 costs from these sources are categorized as to repair parts and labor fuel electricity lubricants tires and ground engaging components bucket teeth tracks etc those preproduction development costs associated with machine use are really just summations of equipment operating costs over the period of time needed to excavate the development openings when the mine is in production daily equipment use varies according to the mine development schedule so estimates should be based on a representative snapshot or series of snapshots labor costs are determined in a similar manner to arrive at daily labor costs estimators need only multiply the number of workers assigned to any one discipline by the number of hours worked per shift and then multiply the result by the associated hourly wage factored for burden wages from mines found throughout north america can be found in labor surveys published by infomine usa the factors shown in table 4 8 1 when applied to average wages for the united states are sometimes used to roughly estimate wages in other parts of the world these factors are based on mandated minimum wages in the respective countries and as a consequence they provide only a rough guideline and should be used with some caution as an evaluation progresses estimators should attempt to gather actual salary and wage data for the region in which the project is located a case can be made that that labor efficiency is proportional to wage rates so that more people are required to achieve the same result in lower wage environments consequently lower wage rates rarely result in proportionally lower operating costs wages must be factored for the additional expenses incurred by the employer for each employee these expenses commonly referred to as burden include contributions to social security taxes worker s compensation and unemployment insurance retirement plans and medical benefit packages additionally evaluators must factor either the wages or the work force to account for the expenses associated with vacation and sick leave shift differential allowances and overtime pay publications available from infomine usa contain extensive details of the costs of these benefits at more than 300 active operations estimators calculate costs for salaried workers in a manner similar to those that the

y use for hourly workers and the sources for salaries are the same as those for wages finally evaluators calculate supply costs by multiplying daily consumption rates by the prices of the consumables these are typically gleaned from individual vendors or from mining cost service infomine usa 2009b as with those associated with equipment operation and labor the expenses associated with supply consumption contribute both to preproduction development and to operating costs estimators need to tally the costs of items such as pipe rail ventilation tubing electric cable rock bolts and shotcrete for each development opening both before and during production along with expenses of drill bits explosives caps and fuses up to this point costs have been estimated in terms of dollars per day the utility of this approach now becomes apparent operating costs are most often reported in terms of dollars per ton of ore and capital costs are typically reported as annual expenditures to report operating costs in the appropriate terms evaluators need only divide the sum of the daily operating costs by the total amount of ore mined each day for capital costs evaluators can simply multiply the daily costs for a specific task by the number of days it takes to complete that task for instance the number of days needed to complete an adit or if the task takes more than a year to complete by the number of days spent on the task each year operating costs typically include a miscellaneous allowance for expenses too small or too numerous to list separately or for expenses associated with unscheduled and unanticipated tasks evaluators sometimes account for such uncertainties by always faulting to the generous side when they calculate each cost estimation parameter however it is preferable to include and list the allowance as one separate value so that those who rely on the estimate can judge its impact for themselves capital costs should include a contingency fund as opposed to the function of the miscellaneous allowance that was included with the operating costs the contingency fund is an actual expense that represents an account set aside for any additional unforeseen costs associated with unanticipated geologic circumstances or engineering conditions the contingency fund is not in place to cover inadequacies in the cost estimate or failings in the mine design but the amount of the fund is typically proportional to the amount of engineering that has gone into the project the money is almost always spent evaluators also need to account for several other expenses in the capital cost tabulation these include costs associated with efforts expended on project feasibility engineering planning construction management administration accounting and legal services for lack of better information estimators commonly factor values for these from the overall equipment purchase plus preproduction development capital cost a variety of

sources report an equivalent variety of factor values but some of the more commonly used factors include the following feasibility engineering and planning approximately 4 to 8 construction supervision and project management approximately 8 to 10 administration accounting permitting and legal services approximately 8 to 14 as an alternative evaluators can base these values on estimates of the time spent on each in conjunction with the salaries of the suitable personnel because most of the expenses are attributable to their work along with the associated office overhead however many of these preproduction tasks are often outsourced and if such is the case the associated expenses should be adjusted accordingly to permit a mine engineers are typically required to submit the results of much of the work that they undertake during the feasibility engineering and planning process to the appropriate permitting agencies estimators are cautioned not to include these expenses twice in their evaluations once as part of the feasibility engineering and planning cost and again as part of the permitting cost economic evaluation to determine the economic viability of a proposed mine evaluators must compare estimated costs to anticipated revenues under the economic conditions linked to the project taxes royalties financing etc as mentioned previously costs are categorized as either capital or operating so that they may receive the appropriate treatment in an after tax analysis operating costs are those that can be directly expensed against revenues as they accrue and include funds that an organization spends operating the equipment purchasing supplies and paying wages and salaries capital costs are those that cannot be fully expensed in the year incurred and include items such as the following exploration property acquisition engineering and construction management mine and mill equipment purchase infrastructure preproduction development buildings contingency fund working capital postproduction reclamation estimators categorize operating costs in several ways production oriented evaluators are typically most comfortable with results that reflect costs in terms of dollars per unit of development e g dollars per meter of drift or dollars per unit of production e g dollars per metric ton mined because operators primarily write checks to the supply vendors the equipment manufacturers or the workers wages and salaries many evaluators prefer to see costs broken down accordingly the choice is really just a matter of preference tempered with intended use because most early stage economic evaluations intend only to estimate overall operating costs the breakdown is not critical only the results the process of an after tax discounted cash flow economic evaluation is beyond the scope of this discussion reliable results are based on many factors in addition to theestimated costs proj

ect revenues for instance are not simply the product of the commodity price the production rate and the resource grade the recovered grade for instance must be factored for losses and dilution at the mine and for concentration inefficiencies at the mill charges that the operator must pay for smelting and refining must be considered as must penalties for deleterious minerals federal and state income taxes as well as sales property and severance taxes reduce anticipated revenues and if operators rely on external financing to back their project or if royalties must be paid to partners property owners or other entities then project economics are further diminished in closing one should keep the estimate in perspective there is no way to exactly predict the costs of a proposed mine and all evaluators know that their estimate will ultimately be proven wrong however evaluators must do their best to minimize the extent to which their estimated values differ from the actual project costs underground mine cost models appendices 4 8a through 4 8c present three cost models that evaluators can use to make preliminary order of magnitude estimates for projects for which there is limited deposit information these models are based on theoretical engineering parameters and do not represent any specific mine they include the following techniques room and pillar mining block cave mining and mechanized cut and fill mining engineers do not rely on models to make significant economic decisions a cost model no matter how carefully the estimator prepares it is only a representation of a hypothetical set of resource parameters and cannot be expected to represent costs for a specific deposit with the degree of reliability necessary for investment models can however be quite useful as comparative tools and evaluators often rely on them to establish cutoff grades for preliminary reserve estimates the figures in the appendices are idealized sketches of the stope layouts for each model model construction the models presented in appendices 4 8a through 4 8c were developed by evaluating sets of hypothetical resource parameters using standard engineering based cost estimating techniques such as those described in the preceding paragraphs to approximate capital and operating costs for underground mine designs based on specific deposit parameters some of the selected salary wage and supply costs on which the program relies are listed in tables 4 8 2 through 4 8 4 these are the most recent values from mining cost service infomine usa 2009b cost estimates for the modeled projects list all of the labor material supply and equipment operating expenses accrued at the mine site including those associated with supervision administration and on site project management also listed are the costs of purchasing and if necessary installing all of the necessary machinery as well as those associated with preproduction development work a

nd constructing the surface facilities costs not included in the estimates are as follows exploration off site roads power lines or railroads taxes except sales tax depreciation off site product transport overtime labor costs milling smelting and refining costs permitting home office overhead insurance town site construction and operation incentive bonus premiums sales expenses interest expense each modeled mine includes at least two routes of access to the deposit for mine models that produce less than 4 000 t d through a single shaft a secondary access raise provides emergency egress and completes the ventilation circuit in the models ore and waste rock densities are 0 367 and 0 401 m3 t respectively ore swells to 155 of its in place volume on excavation and waste swells to 145 of its inplace volume rock quality designations and compressive strengths vary from one model to the next values for several methods from a variety of mines are listed in table 4 8 5 preproduction development work blocks out enough ore to initiate operations at the design production rate and the level of production development work is designed to maintain that rate throughout the life of the mine all shop office worker changehouse warehouse and mine plant buildings are constructed on the surface working capital allows for 2 months of project operation and a sales tax rate of 6 75 is applied to all equipment and nonfuel supply purchases capital costs do not include the expenditures associated with outside contractors infrastructure home office overhead insurance or project startup except working capital costs are in late 2008 and early 2009 dollars unit costs wages and salaries used in the models represent u s national averages as reported in u s metal and industrial mineral mine salaries wages and benefits 2009 survey results salzer 2009 in keeping with the results of that survey lower wages and salaries are used for the smaller mines and higher wages and salaries are used for larger mines in the models the cutoff point between small and large mines is set at 100 employees equipment and supply prices are for the most part taken from mining cost service infomine usa 2009b in the models the salaries shown in table 4 8 2 are adjusted upward to account for a 38 0 burden rate at the small mines and a 44 0 burden rate at the larger mines hourly wages used in the cost models are shown in table 4 8 3 the wages shown in table 4 8 3 are adjusted upward to account for a 38 0 burden rate at the small mines and a 44 0 burden rate at the larger mines it is obvious that costs vary from one mine to the next so although it is of interest to know the costs associated with surface mines in general terms it is also important to understand how to estimate the costs of a proposed operation in a way that considers the unique development and operational parameters and subsequently costs of each depos

it although focusing primarily on how to estimate costs this chapter also includes general operating expenses for typical surface mine configurations there are probably as many ways to estimate mining costs as there are cost estimators because of the lack of a standardized approach evaluators are left to estimate costs as best they can so almost everyone uses a slightly different method a standardized method that suits every situation would be extremely difficult to develop given that each proposed mine is unique and conditions can be so variable although no such approach exists many well documented methods are available for example there are the tried andtrue broad brush approaches one of which is the parametric method where costs are derived from general algorithms or curves of the following form cost x parameter y the parameter in these algorithms can be almost anything but most often it is the production rate the x and y values are derived through statistical evaluations of known or estimated cost data the u s bureau of mines cost estimating system also known as ces usbm 1987 can be considered a parametric approach as can methods developed by o hara 1980 and mular 1982 another example of a broad brush method is the factored approach usually with this technique one primary cost such as the cost of the purchased equipment is subjected to a series of factors to estimate all the other pertinent costs of the project vilbrandt and dryden 1959 this method has fallen out of general use because it is in light of subsequent approaches considered too general evaluators also commonly rely on a comparative approach with this method estimators examine costs at similar projects and make adjustments often through the use of scaling factors schumacher and stebbins 1995 to account for differences in operating parameters this may be the most comforting of the broad brush approaches but it can also be the most misleading conditions simply vary too much from one project to the next to rely too heavily on comparative costs if conditions were the same at every deposit then assigning costs from a past or similar project would be acceptable and the approach would be widely used but it is the differences in the operating parameters from one project to the next that dictate the differences in costs so these must be fully considered cost models are a form of the comparative approach these consist of a compilation of cost estimates along with the parameters on which those estimates are based evaluators find the example from within the compilation that most closely resembles their project and they then use the costs associated with the example as an indication of the costs at their project example cost models for typical surface mine configurations can be found in appendix 4 9a significant effort went into the derivation of the specific variations of the aforementioned methods and each represents an invalua

ble source of useful reliable information in particular the ces curves usbm 1987 enable evaluators to estimate costs for a multitude of mining and mineral processing activities for which no other source exists but arguably the concern with each of these approaches is the lack of transparent detail evaluators are left to wonder if results truly represent their project even though broad brush methods are often used because much of the information needed for more detailed analyses is difficult to obtain evaluators still continually strive for more verifiable and hence reliable results in the past the broad brush approaches also maintained their popularity in part because more detailed analyses were time consuming over the past 20 years however things have changed most evaluators now use a more detailed engineeringbased approach to estimating costs at almost every stage of project evaluation two events have led to this eventuality the first was the development publication and distribution of mining cost service infomine usa 2009b along with an increase in the availability of information similar to that contained in mining cost service through the internet this annually updated document is a comprehensive compilation of current mine and mineral processing cost information the second event was an improvement in spreadsheet and application based calculation modeling capabilities which enabled evaluators to handle the significant increase in the amount of work associated with engineering based estimates in a timely manner evaluators now conduct engineeringbased estimates in time frames previously achievable only when they used the broad brush approaches engineering based itemized cost estimating the method detailed in the next few paragraphs is best described as an engineering based abbreviated itemized approach it consists of three major steps along with a highly variable number of minor steps in the first step estimators design a mine to the maximum extent possible given the available information for a deposit that can be mined using surface techniques even a general pit outline an overall depth and a delineation of the routes to the processing plant and the waste stockpiles provide a great deal of information pertinent to the cost estimate in the next step evaluators estimate or calculate all the parameters associated with the things that cost money the workers the equipment fleet and the consumable supplies this step is where an estimator expends the most effort although the first design step previously outlined is the most important in achieving reliable results the final step is the simplest thanks to publications such as mining cost service evaluators need only apply known unit costs for labor equipment operation and supplies to the projected and calculated development and operating parameters to arrive at estimates of the operating costs in addition to estimates of many of the preproduction

development costs they then need to apply equipment purchase prices along with the costs of some common mine facilities to the previously determined parameters to arrive at the primary components of a capital cost estimate the advantages of the engineering based itemized approach are many it can be applied at almost any stage of a project evaluation from the initial phases when information is scarce to the final stages when almost all pertinent resource and project characteristics have been established it is reliable in that it concerns itself almost exclusively with parameters specific to one deposit it lends itself well to computerization because so much of the work involves simple calculations albeit a lot of them that are easily encoded on a spreadsheet or a windowsbased application it is easily adjusted and updated as more information becomes available as such the reliability of the estimate increases as the information base expands and when the evaluation stage is complete the final computerized product is in essence a dynamic cost model that engineers can use to examine operational alternatives throughout the life of the mine traditionally and logically evaluators have kept the level of detail in their cost estimates comparable with the amount of information available for the deposit unfortunately it is sometimes tempting to reduce the level of detail in an effort to reduce the amount of time spent on the estimate ignoring detail by procedures such as averaging site parameters or combining cost components can reduce the representativeness of the estimate for instance if haul distances and gradients for individual haul segments can be gleaned from maps and plans the cycle time associated with the haul may be significantly different than the cycle time for a more convenient but less reliable overall distance and average gradient over the entire distance example 4 presented later in this chapter helps to illustrate this point just as significantly combined cost values such as those presented for equipment operation in various publications can also lead to estimates that are not fully representative if such costs are broken down into individual components i e fuel lubricants repair parts tires and wear parts then each component can be adjusted individually to suit conditions for instance in a situation where a mobile loader is used to collect extremely abrasive rock the evaluator might adjust the tire and wear part consumption rates upward if these components were not treated separately the evaluator might simply adjust the entire composited operating cost upward the significance of avoiding such an approach is this if you increase the tire consumption rate by 100 i e multiply the tire operating cost by 2 and the tire cost is initially 10 of the overall operating cost then the impact on the overall machine operating cost is minimal as would be any error in the evaluator s assumption

of the increase and because equipment operating costs may only represent 25 of the overall operating cost the impact of any error would be even less in essence a 100 error in a cost component that comprises only 2 5 of the overall cost is much less significant than a similar error in a cost component that comprises 25 of the overall cost getting started often where to start is the question it is sometimes a difficult question to answer when an estimator is trying to figure out how much a deposit will cost to mine however when evaluators begin the process of approximating the costs of a mining project they soon notice a synergy as one parameter is determined the value of another is often defined for instance as the number of trucks needed to haul the ore is determined the number of drivers required to operate the trucks and the number of mechanics needed to maintain them are also determined an evaluator can then use those values to begin the process of estimating the sizes of the shop the parking lot the living quarters if needed and the workers changehouse drill and blast to estimate the costs of drilling and blasting engineers can glean a great deal of information from just a powder factor such a factor which is most often reported in terms of kilograms of explosive per metric ton blasted of course differs from one project to the next and is typically determined through experimentation observation and adjustment over time at an active operation consequently the value will not be known ahead of time but reported powder factor values are plentiful in books such as this handbook in case studies contained in periodicals and in publications such as the mining source book scales 2009 a powder factor from a mine in rock similar to that of a proposed project should supply an initial value that is within reason from this one value engineers can of course estimate the cost of explosives in terms of dollars per metric ton of ore but in addition they can also estimate how much in terms of meters to drill each day which in turn provides the number of blastholes that must be drilled each day and that value in turn furnishes the number of caps and boosters consumed each day with the daily drilling requirements in hand estimators can approximate values for daily drill use in terms of hours per day drill bit and steel consumption and with all this previous information they can proceed to gauge the required number of drillers and blasters all of this is a lot to derive from just a powder factor and it is important to remember that for an early stage cost estimate precise values are not necessary nor can they be expected reasonable representative values are required but highly precise values are simply not obtainable at the early stages of a cost estimate unless the information needed for such precision exists to illustrate the process example 1 works from a powder factor to estimate consumption ra

tes and subsequently costs for explosives caps detonation cord and drill bits and steel from there estimated values are further used to suggest drill use in terms of hours per day as well as labor requirementexcavate and haul estimators find that most of the expense of any surface mine is attributable to excavating the rock loading it into some sort of conveyance hauling it somewhere either a mineral processing plant or a stockpile and then dumping it consequently a representative estimate hinges on the reliability of the excavating and hauling costs as with the cost estimates of all the other surface mining tasks the basis for the costs of excavating and hauling begins with the design it is crucial to know the routes over which the ore and waste will be hauled the more that is known about these routes the more reliable the estimates will be distances and gradients are the key components and while average gradients over total haul distances can be used much more reliable results are achieved if the routes are split into segments at each significant change in gradient the importance of carefully defining the distances and gradients of each segment increases with the stripping ratio evaluators find that large projects with high stripping ratios can become in essence waste bound in that the space needed to stack and store waste is at a premium at such deposits operating costs are more sensitive to waste haul distances and gradients than to any other factor to estimate excavating and hauling costs evaluators must first determine cycle times for both the excavators and the haul trucks evaluators use these cycle times in conjunction with respective machine capacities to gauge the size of the required fleet and to eventually estimate operating costs and purchase prices if the purpose of an evaluation is to estimate the average costs of production for the project then the haul profiles should be defined at a point halfway through production in other words they should be based on the pit profile at that point in time when about half the resource has been extracted when engineers structure the cost estimating process on a spreadsheet or through a windows application or any number of other computerized approaches it is entirely possible for them to estimate the costs associated with haul profiles from any bench in fact from any point on any bench in the pit this is of course pertinent when an evaluator is optimizing a resource with software that asks for production costs from various benches as part of the optimization process cycle times for excavators are for the most part fixed and related to machine size wheel loaders are the exception in that they are sometimes called on to travel a short distance from the active face to the loading point most tracked excavators simply pivot after they collect a load of broken rock to transfer that load to the truck truck cycle times are more complicated although so

me of the time components are fixed spot load dump and turn travel times typically represent the largest component of a truck s cycle it is also the component that typically has the greatest impact in distinguishing costs at one project from those at another engineers attempt to achieve the following goals as they design the excavator and hauler segments of their mine plan three to six loader cycles should completely fill the truck bed loader bucket capacities should be selected so that whatever the number of cycles the truck is full or close to full after loading is complete for instance a 7 0 m3 bucket could be used to fill a 21 0 m3 capacity truck but it would be inefficient if used to fill a 17 0 m3 capacity truck two loads would not fill the 17 0 m3 capacity truck completely but three loads would overfill it the number of trucks and the number of loaders should be determined to minimize both the amount of time that any loader must wait for a truck and the amount of time that any truck must wait in a queue to be loaded to meet these goals engineers rely on a multistep process first they estimate the loader cycle time and use it in conjunction with the loader s bucket capacity to determine the number needed to meet production goals this first step is straighto produce a complete cost estimate much work remains for the evaluator even after the drilling blasting excavating and hauling costs have been determined however these previously determined costs along with the parameters derived during the estimation process do provide a basis for estimates of the remaining costs to begin with the costs associated with many of the machines typically found at any surface mine have yet to be estimated for most such projects costs for bulldozers graders dust suppressant tankers equipment maintenance trucks pumps lighting plants personnel movers and in some cases generators crushers and conveyors may all need to be estimated and included for each of these machine types the determining factors that provide the basis for the estimated costs are as with the drills excavators and haulers the capacity of the machine and how many hours it must operate each day consequently the techniques that evaluators rely on to estimate the capacity and daily use parameters for each are similar to those they use to gauge the same parameters for the drills excavators and haulers for instance at almost every surface mine a fleet of bulldozers manages blasted rock at the working faces and dumped waste rock at the stockpiles in addition to performing a host of other tasks the process that estimators use to determine the number and operating requirements of these machines is rarely as straightforward as the process that they use to determine the excavator and loader needs but it is still based on a very similar approach at all but the smallest operations bulldozers work continuously at each dump site they also o

ften work at each active face moving scattered broken rock to the excavator the size requirements for these machines are based on the amount of material that they handle each shift and the distance that the material must be moved specifically each blade load carries with it a volume that will be moved over a distance at a speed typically specified in the manufacturer s documentation for the purpose of project evaluation costs are typically categorized as either operating or capital as opposed to fixed or variable so that they can be subjected to after tax discounted cash flow analyses in short operating costs are those that can be fully expensed in the year incurred the expenses of the consumables including those associated with equipment operation wages and salaries are typically all considered operating costs and are most often estimated either in terms of dollars per metric ton of ore or dollars per year capital costs are those that cannot be fully expensed in the year incurred and include items such as the following mine and mill equipment purchase development engineering and construction management infrastructure working capital postproduction reclamation preproduction stripping property acquisition exploration buildings contingency fund although this chapter deals primarily with costs and cost estimation it is worth mentioning that from an after tax economic viability standpoint it is best to minimize the preproduction capital expenses and incur them as close to startup as possible because of the time value of money capital expenses accrued later in the operation have a lesser impact on the overall project net present value so during the preliminary mine design process evaluators find that it is worth the effort to structure the project in a way that expedites production cost components while expenses at a mine can be categorized as either capital costs or operating costs both are comprised almost entirely of labor supply and equipment components whether building a processing plant constructing a tailings impoundment or mining an ore deposit evaluators find that most of the money spent on the project goes to either the workers laborers skilled tradesmen equipment operators supervisors technicians managers etc the supply vendors to purchase wood drill bits concrete steel explosives tires diesel etc or to equipment manufacturers to purchase machines or buy parts labor wages and salaries and the burdens associated with each very often represent the largest expenditures at any mineraldevelopment project in fact wages benefits mandated employment taxes and bonuses can sometimes account for more than half of the total operating costs depending on the size of the mine labor costs can account for anywhere from 15 to 60 of the total operating costs these costs can escalate if the mine is situated in a remote area without a local source of skilled labor wag

es also tend to be one of the more variable components of an evaluation project location has a significant impact and evaluators are urged to examine wages on a regional level to properly account for the associated expenses in their evaluation wages and salaries for miners in several countries are tracked and reported in publications such as u s metal and industrial mineral mine salaries wages and benefits 2009 survey results salzer 2009 base wages are loaded with mandated employment taxes including social security medicare unemployment taxes and workers compensation taxes other items add to the burden factor such as shift differentials overtime medical dental and vision benefits retirement plans short and long term disability insurance life insurance accidental death and dismemberment insurance sick leave vacation and holiday pay and other benefits to retain employees companies often use creative benefits such as paid tuition transportation to remote mine sites attendance bonuses safety bonuses family and individual assistance plans and paid fitness club memberships many mines pay their production miners a bonus based on meeting development or production goals these bonus systems are sometimes modified to include safety ground conditions and other factors safety violations can reduce or even eliminate a production bonus other criteria sometimes used to calculate bonuses are individual performance safety performance commodity price profit recovery ore grade production and cost savings to calculate bonuses supplies supply prices are less volatile than wages and salaries but they still vary from one region to the next and from one vendor to another while it is always preferable to obtain local prices from established vendors it is often impractical to do so during the early stages of project evaluation mining cost service infomine 2009b provides an extensive array of supply costs that are reliable for early stage feasibility work in the evaluation process some supply costs are commonly reported as equipment operating costs because their consumption rates are directly tied to machine use diesel fuel gasoline electricity tires and lubricants all fall into this category and as demonstrated earlier many of the project s labor requirements and subsequent costs are also directly dictated by daily machine use in addition to individual equipment operator requirements mechanic electrician machinist and equipment maintenance was with the supply prices equipment purchase prices are typically obtained from vendors however in the early stages of an evaluation it is even more difficult to obtain these values than it is the supply costs because the necessary machines can only be specified in the most general terms infomine usa s mining cost service infomine 2009b and mine and mill equipment costs an estimator s guide infomine usa 2009a also contain extensive purchase price lists for

machines commonly used at surface mines for early stage feasibility work equipment prices are usually based on list prices as suggested by the manufacturers with no discounts assumed and no options added early in the evaluation process most evaluators specify new machines for all production related project requirements and purchase prices reflect this if anticipated use is minimal estimators may specify previously owned used machines for some of the secondary support equipment such as water tankers and road graders if previously owned machines are relied on for production work the equipment productivity and availability and the associated operating costs should be adjusted accordingly in anticipation of increased maintenance and repair requirements cost models when evaluators have limited deposit information they can use mine models for order of magnitude estimates in addition models can be used to provide insight into the nature of mining costs in general the impact of changes in operating parameters can be easily understood when presented in a format that compares costs associated with one configuration directly to those associated with another unexpected changes dictated by increases in stripping ratios at the larger operations and also note the ratio of labor costs to equipment operating costs as production rates increase these models are theoretical and are not representative of any existing mine note the pit and haul parameters which provide the basis for each design these should be one of the key points of comparison if the models are to be used to provide estimates for any proposed operation costs associated with each model account for all pertinent labor material supply and equipment operating expenses accrued at the mine site costs for supervision administration and on site project management are all included expenses associated with preproduction development equipment purchase and installation and building and facility construction are also included in these models costs for the following operations and facilities are considered ore and waste drilling blasting and excavation ore haul from the active face to the mill site overburden and waste haul from the active face to the dumpsite constructing and operating the facilities required for equipment maintenance and repair electricity and fuel distribution drainage explosives storage and sanitation constructing a mine office a warehouse and a worker changehouse plus all associated site work the relative merits of surface and underground mining are widely discussed and frequently debated some deposits can be mined entirely with surface methods while others can only be worked underground with all other conditions equal surface mining is normally regarded as preferable because of lower development costs quicker start up time and lower accident rates generally associated with surface mining when choosing between surface and unde

rground methods some of the factors that must be considered include size shape and depth of the deposit geologic structure and geomechanical conditions productivities and machinery capacities availability of experienced work force capital requirements and operating costs ore recoveries and revenues safety and injuries environmental impacts during and after mining reclamation and restoration requirements and costs and societal and cultural expectations some deposits may reasonably be mined entirely by surface methods in general such deposits are close to the surface and have a relatively uniform geology similarly some deposits can only be mined economically by underground methods these deposits are usually deeper with geological and mineralogical characteristics that require more selective ore extraction finally other deposits are best mined initially as open pits with production shifting to an underground method as deeper portions of the ore body are extracted an example of each type of deposit follows in suitable deposits surface mining is more productive more economic and safer for workers however changes in environmental regulations and societal expectations may lead to fewer large open pit mines particularly if operators are required to backfill open pits and recontour waste dumps these conditions may result in the development of small high grade deposits by very shallow open pits or in the development of high grade underground mines in place of large open pit mines where applicable large low grade deposits may be mined by in situ methods hitzman 2005 in some cases especially in built up areas it has become almost impossible to obtain permits for new surface mines this is the case for producers of crushed stone and dimension stone in large metropolitan areas in many developed countries for this reason several underground quarries have recently begun operating in the united states and many more are in the planning stages surface and underground mine examples in some cases the choice of surface or underground is obvious one such example is the north antelope rochelle mine in wyoming united states owned by peabody energy the north antelope rochelle mine shipped 88 7 mt of compliance coal in 2008 and has produced more than 1 000 mt since the mine began in 1983 it is the largest coal mining operation in the united states remaining coal reserves dedicated to the mine cover nearly 8 800 hectares with about 1 200 mt of recoverable coal the coal seam ranges from 18 to 25 m thick and lies from 15 to 105 m below the surface the complex employs a large fleet of big equipment listed in table 6 1 1 the three draglines have bucket capacities of 84 76 5 and 65 m3 respectively the key to success is high volume and low unit costs similarly in the case of the henderson mine underground mining by panel caving was the most logical choice the henderson mine located in colorado u

nited states is owned by freeport mcmoran copper and gold inc a cross section of the henderson ore body is shown in figure 6 1 1 although the ore body is relatively large it is also quite deep about 1 040 m below the top of red mountain development of an open pit mine would have required removal of a large amount of overburden before the ore body was exposed this would have required construction of roads power lines and other infrastructure to the top of red mountain at 3 751 m above sea level thus development would have been extremely expensive with no initial production to support development costs the henderson mine was developed as shown in figure 6 1 2 finally the northparkes mine in queensland australia provides an example of a mine that began as an open pit and is now an underground operation as shown in figure 6 1 3 the northparkes ore body is a narrow porphyry 200 300 m across and about 900 m in height beginning in the late 1970s northparkes was mined with two open pits each about 150 m deep because the ore body has such a small cross section the stripping ratio increased rapidly and in 1993 development was begun for underground mining by block caving as shown in figure 6 1 4 with these examples in mind it is worthwhile to consider some specific differences between surface and underground mining production much more material is produced by surface than by underground mining this is shown in table 6 1 2 which gives recent data for the united states it is apparent from table 6 1 3 that sand gravel and stone products represent more than 90 of the material produced in surface mines each year the fraction of mined material produced by underground methods in the united states has decreased in recent years as shown in table 6 1 3 this results from the decrease in the fraction of coal produced underground the fractions for metal and nonmetal minerals vary over the period mine size in terms of daily production tonnage surface mines are almost always larger than underground mines producing the same commodity this is partially true because open pit mines must mine much more waste rock and therefore have much more dilution of the in situ mineral whereas many of the underground methods can mine the same mineral much more selectively with less dilution and therefore fewer metric tons table 6 1 4 shows approximate daily production rates for selected large surface mines table 6 1 5 shows similar data for large underground mines these tables show the predominance of surface methods for large high tonnage operations worldwide tables 6 1 4 and 6 1 5 are not intended to be complete but are included to provide an indication of the respective numbers and sizes of larger surface and underground operations productivity when productivity is measured in metric tons mined per worker hour surface mines are almost always more productive table 6 1 6 shows data for coal mining in the united states during 2006 a

nd 2007 productivity in surface mines was more than three times that in underground mines however when choosing a mining method it is important to go beyond a simple consideration of metric tons per workerhour for example in a gold deposit it may be more meaningful to examine grams or ounces of gold produced per worker hour in many gold mining districts comparing the productivity of the surface mines and underground mines in this way shows much more comparable results safety the mining industry throughout the world continues to reduce the incidence of accidents and fatalities the underground mining environment is recognized as being more hazardous than the surface table 6 1 7 shows the incidence rates per 200 000 hours for all accidents in the united states during the years 2003 2007 table 6 1 8 shows the incidence rates for fatal injuries these rates are higher in all cases for underground mining and notably higher for underground coal mining development development for surface mining of coal and other bedded minerals involves the removing of cover layers of soil and rock to expose the coal surface mining is used when the coal seam is relatively close to the surface usually within 60 m the time between overburden removal and the mining of the product mineral should be as short as possible to optimize overall cash flow however for larger deposits covered by large amounts of overburden and waste the amount of pre stripping will also be large leading to high preproduction development costs the time required for prestripping can range from 2 to 6 years thus interest costs during development will be high and will represent a significant portion of the pre mining capital requirement before mining can start when an ore body is steeply dipping and at or near the surface open pit mining can start with a small amount of stripping however as mining of such a deposit progresses increasing amounts of waste rock must be removed this must often be done many years before mining of the corresponding amount of ore at deeper levels can take place thus the ultimate pit limits must be projected early in the mine planning process and the investment cost for waste rock removal in advance of mining must be included in the economic evaluation waste rock stripping should be delayed as long as possible to avoid high interest cost for all the money spent in waste stripping activities the increasing cost of stripping at greater depths is one of the major factors in deciding when to transition from surface to underground mining of a given deposit in an underground mine a significant amount of infrastructure must be installed before mining begins this will include shafts hoists ventilation fans underground shops travel ways for workers and machinery ore storage bins underground crushers and so forth this requires detailed long range planning from the very beginning so that the requirements of future workings at deeper levels can

be accommodated a large capital investment is often necessary before production can start underground mining methods require a more careful design and planning process because it is difficult to make changes in a design after the infrastructure has been installed and the equipment purchased this condition is often exacerbated when variables such as ore grade mine water make and ground control conditions change or are different than expected it is very important that the underground mine design and the machinery capacities are properly chosen from the beginning for all of these reasons it is often prudent to develop a small test mine to accurately determine many of the unknown mine characteristics a test mine and a properly conducted feasibility study will minimize these risks the development of a large underground mine can take as many as 5 to 10 years interest costs during development will therefore be high and may comprise 30 to 40 of the premining capital requirement before mining can start cost comparisons estimates of capital and operating costs for surface and underground mines of various sizes and configurations are compiled regularly and in considerable detail by infomine inc those estimates are provided to customers as the mining cost service and can be purchased in printed or electronic form or accessed on line the cost estimates do not include permitting environmental analysis reclamation or closure costs figures 6 1 5 and 6 1 6 summarize the cost estimates for surface mines the mining cost service also provides estimates of capital and operating costs for underground mining the data are more extensive with estimates for eight mining methods and shaft and adit access for each figures 6 1 7 and 6 1 8 summarize selected data while surface mining methods are relatively simple and uniformly applied there are many underground mining methods and application of any given method will vary from mine to mine thus it is much more difficult to accurately summarize costs for underground mining methods nonetheless figures 6 1 5 through 6 1 8 show the following trends for small mines capital and operating costs per metric ton of ore produced are lower for surface methods of course dilution and ore grade must also be considered in a full economic analysis for large tonnage production capital and operating costs may be higher for surface mines depending on stripping ratio in these cases a dual feasibility study must be performed comparing the openpit option to the best underground mining option in all cases capital costs increase and operating costs decrease with increasing production tonnage environmental and closure requirements surface mines create a much larger footprint than underground mines in the united states surface coal mines are required to backfill mine excavations and recontour and revegetate waste piles this is not the case for metal but there are strong indications that the situati

on is changing indeed in most countries society no longer looks favorably on large abandoned excavations and the mandated costs of reclamation and closure for large surface mines are likely to increase a permit for construction of a new surface mine or expansion of an existing surface mine cannot be obtained in some areas in these cases underground mining should be examined costs of environmental compliance reclamation and closure are seldom published however in 1999 mudder and harvey reported that closure costs for u s surface mines ranged from us 1 236 to us 3 707 per disturbed hectare with coal mine costs on the higher end of the range at us 2 471 to us 3 707 per hectare costs for metal mine sites were lower yet they were much higher in cases where extensive water management and acid rock encapsulation were required in 1996 homestake mining company reported average company wide reclamation costs of us 3 361 per hectare between 1980 and 1992 136 abandoned coal mine sites in pennsylvania were reclaimed at a cost of about us 2 348 per hectare bogovich 1992 in 2004 wilson and dyhr estimated environmental and closure costs as a percentage of total operating costs for medium sized mines with on site processing and tailings disposal those estimates are summarized in figure 6 1 9 where the higher costs associated with surface mining are clearly shown selection of a mining method based on this brief introduction it may appear that surface mining is preferable to underground methods particularly in regard to productivity and worker safety however as has been pointed out selection of the best mining method for any deposit requires analysis of many factors besides the simple productivity in metric tons of ore per worker hour the following subsections discuss in detail the procedures for selecting a mining method and include factors that influence the choice between surface and underground mining location of the deposit in some cases a mineral deposit may be located in a place where a large surface mine is simply unacceptable the case of stone and other construction materials already mentioned briefly is an excellent example these materials have a relatively low value and are used in large quantities so it is important that they be mined as close as possible to the locations where they will be used those places are almost always heavily built up areas for example the amount of concrete used in manhattan every 18 months 3 33 million m3 is about the same amount as was used in the construction of the hoover dam on the border between arizona and nevada in the united states owen 2003 in such areas it is difficult if not impossible to expand an existing stone quarry let alone open a new one and the production of stone increasingly comes from underground quarries of course underground mining methods can also have adverse effects when operated under built up areas surface subsidence and mine water release must

both be monitored and controlled other factors also enter into choosing the location for a mineral deposit including processing requirements political and social conditions and work force availability which are discussed in the following sections another factor environmental and permitting requirements is not discussed in this chapter geology of the deposit three aspects of a deposit s geology relate to the choice of surface or underground mining the intrinsic value or grade the morphology and the structure a material with a higher intrinsic value will support a more expensive mining method for example jim walter resources mines high quality metallurgical coal at its blue creek mine in alabama united states under very difficult conditions that include spontaneous combustion deep cover 450 to 730 m and high methane levels howell et al 1991 a lower quality coal would not support the high costs of mining in this geological setting similarly an unusual narrowvein gold deposit where the gold occurs in very high grade but sporadic pockets supports a labor intensive underground mining method original sixteen to one mine 2009 in general deposits with lower intrinsic value or grade are more amenable to surface mining methods when other conditions permit deposit morphology including shape extent and depth is also important the economics of most surface mining methods and some underground methods are based on high production volume and low unit costs and use of equipment that has high capital costs these require large deposits with relatively uniform grade and few irregularities in shape or extent deposits that meet these criteria can often be mined profitably even when the ore grade or product value is relatively low good examples of such surface mines are large coal mines in northeastern wyoming as described previously the large porphyry copper mines such as bingham canyon in utah united states and chuquicamata in chile and the large low grade gold mines such as round mountain and goldstrike in nevada similarly coal deposits that can be mined by the underground longwall method must have large areas of coal with relatively uniform thickness to allow the development and production of large panels that will support the costs of development and purchase of equipment the depth of a deposit also influences the surface versus underground decision the depth of the blue creek mine requires the use of barrier pillars between longwall panels at a cost that could probably not be supported by a lower value coal in other cases metal deposits are often mined initially by the open pit method but switch to an underground method when the costs of removing overburden become too high this has happened for example at kiruna in sweden northparkes in australia and palabora in south africa finally the geologic structure of a deposit must be considered it is more difficult to generalize about this factor but a

good example is the homestake deposit in south dakota united states george hearst who consolidated the claims and put them into production is reported to have said to his partners here s to low grade ore and plenty of it smith 2003 during its 125 years of operation the homestake mine produced almost 1 2 million kg of gold and 0 3 million kg of silver of course open pit mining was unknown when operations began at homestake but lacking other information one might conclude that this large low grade deposit was an ideal candidate for that method however the deposit was highly folded and faulted and required selective mining to extract the ore in a manner that could only have been done by underground methods processing requirements the processing required to produce an economic product also influences the choice of mining method it may be possible to mine a low grade ore at very low cost using a surface method but the resulting dilution may make processing so expensive that the overall operation is not profitable in such a case more selective mining using an underground method may be used to produce a higher grade ore which is less expensive to concentrate such selective mining can also be used to leave in place portions of the deposit containing impurities or contaminants that can increase reclamation or remediation costs if they enter the process stream it is also important to consider the locations of available processing facilities and the ease with which new facilities can be permitted and built the difficulty of obtaining permits for new operations in built up areas for aggregate pits and stone quarries was described previously the same challenge has been encountered in permitting new coal preparation plants in the eastern united states and in some cases this has been the main factor in deciding how to mine new coal resources political and social conditions political and social conditions can determine not only whether or not a mineral deposit can be mined but also the method by which it is to be mined there may be significant opposition to the large highly visible disturbance that occurs in surface mining making permitting too expensive or impossible in other cases the legal rights to the minerals in an area may be separated from the rights to control the surface in the same area so that accessing and removing the minerals by an underground method is preferable work force in general underground mining methods require a more specialized work force than surface methods workers with experience in operating heavy equipment in agriculture or construction can often transfer their skills for use in surface mining operations but underground mining equipment and processes are significantly different for a deposit in which there is no clear choice between surface and underground mining based on other constraints the presence or absence of a suitably skilled work force can be a deciding factor the purpose of

a classification system for mining methods is to provide an initial guideline for the preliminary selection of a suitable method or methods its significance is great as this choice impinges on all future mine design decisions and in turn on safety economy and the environment the choice of a mining method assumes a previous but cursory knowledge of the methods themselves it also assumes a brief understanding of ground control and of excavating and bulk handling equipment in the formal mine design procedure the choice of mining methods immediately follows geological and geotechnical studies and feeds directly into the crucial milestone diagram where regions of the property are delineated as to prospective mining methods lineberry and adler 1987 this step in turn just precedes the subjective complex and critical layout and sequencing study to develop the proposed classification system adopted here many existing ones both domestic united states and foreign were examined and incorporated to varying degrees the result is deemed more systematic inclusive and understandable than its predecessors i e stoces 1966 subsequent parts of this handbook elaborate on the selection and comparison of mining methods input statement a comprehensive statement has been developed to provide a rapid checklist of the many important input parameters adler and thompson 1987 the three major areas are 1 natural conditions 2 company capabilities and 3 public policy table 6 2 1 those parameters appearing early are generally the most important natural conditions require that a dual thrust be maintained concerning resource potentials and engineering capabilities an additional basic distinction occurs between geography and geology for company capabilities fiscal engineering and management resources must be recognized this includes the scale of investment profitability and personnel skills and experience public policy must be considered particularly as to governmental regulations especially safety health and environmental tax laws and contract status some of the latter input factors are held in abeyance until near the end of the investigation and then considered as modifying factors this organization duplicates but tightens others hartman 1987 spatial description most mineral deposits have been geometrically characterized as to an idealized shape inclination size and depth complex or composite bodies are then composed of these elements ideal shapes are either tabular or massive with chimneys or pipes being subordinated tabular deposits extend at least hundreds of meters feet along two dimensions and substantially less along a minor dimension massive bodies are approximately unidimensional cubic or spherical being at least hundreds of meters feet in three dimensions a modification is recommended later to achieve closure with tabular deposits for tabular deposits the inclination attitude or dip and

thickness are crucial inclinations range from flat to steep table 6 2 2 hamrin 1980 popov 1971 in surface mining the inclination limits the advantageous possibility of being able to cast waste material nearby as opposed to hauling it a distance and then storing it for flat deposits especially when fairly shallow an area can be successively opened up and the waste can then be cast into the previously mined out strips a substantial economic advantage casting in its normal sense is not restricted to the use of rotating excavators broadly it means relatively short distance hauling of waste which can also be done with mobile loaders and or trucks or with mobile bridge conveyors for steeper and deeper deposits stable pit slopes become important table 6 2 3 hartman 1987 popov 1971 where the deposit inclination exceeds that of the stable slope both the hanging wall and footwall must be excavated and the increased waste then handled and placed for both surface and underground mining methods the inclination cutoff values nearly coincide one for pit slopes the other for face bulk handling mechanisms whether mechanical or by gravity while not identical they are close enough to use similar values 20 and 45 see table 6 2 2 the thickness of a tabular deposit is also important table 6 2 4 with reference primarily to underground work popov 1971 when three or more benches are required the deposit tends to be treated as massive primarily in flat underground deposits thickness governs the possible equipment height low profile and in steep ones its narrowness also in underground mining the deposit thickness becomes a support problem especially if effective pillars become so massive that recovery is compromised when the upper limit of any of these concerns is reached e g benching equipment size and pillar bulk closure with massive deposits occurs for all practical purposes pillar size vs recovery can dictate caving except where pillar sizes may be decreased because backfilling is used such as in postpillar cut and fill finally the depth below the ground surface is important table 6 2 5 popov 1971 stefanko 1983 for surface deposits even flat ones this can obviate casting and require increased waste haulage and expanded dump sites for underground mining earth pressures usually increase with depth consequently raising the support needs the ground surface location above a deposit must be clearly identified to evaluate other parameters see input statement section previously correlating deposit types the inclination dip can be roughly related to the deposit type table 6 2 6 rocks can also be related to strength table 6 2 7 hartman 1987 the strength of the deposit and its envelope of country rock can then be related to its type table 6 2 8 for determining pit slopes surface mining and support requirements underground mining these relationships become important some

variations are noted especially for veins and disseminated deposits classifying surface mining methods depth related to inclination the surface mining classification although based on the crucial ability to cast waste material rather than to haul it has other features these are primarily based on the depth of the deposit being a function of its inclination flat seams tend to be shallow and casting is possible steep and massive deposits trend to depth from this a number of relationships result depth related to excavating technique and stripping ratio because of the effects of weathering and stress release excavating becomes more difficult and expensive with depth following a continuum from hydraulic action and scooping through to blasting hartman 1987 as a matter of definition the stripping ratio ratio of waste to mineral usually increases with depth however the relatively inexpensive handling of waste near the surface by casting tends to mitigate this increase permitting higher ratios the use of mobile cross pit high angle conveying allows greater pit depths and along with the mineral value also influences this ratio surface mining classification system based on the foregoing factors a surface mining classification has been developed table 6 2 9 the classification incorporates information dependent on the intrinsic characteristics of the geometry of the deposit quarrying appears to be anomalous because of 1 relatively steeper pit slopes 2 specialized means of excavating and handling and 3 less critical amount of overburden glory hole mining or its equivalent is making a comeback in very deep open pits using inclined hoisting glory hole mining utilizes a single large diameter raise located in the lowest point of the pit down which all blasted material is dumped the bottom of the hole feeds into crushers and a conveying system which transports the material to the surface through a horizontal or inclined drift darling 1989 in contrast to the underground classification the surface one is not formed into a matrix this is because depth and therefore the excavating technique waste handling and stripping ratio are all functionally related to the deposit geometry particularly the seam inclination no preceding classification recognizes this relationship hartman 1987 lewis and clark 1964 morrison and russell 1973 stout 1980 thomas 1973 classifying underground mining methods normally two major independent parameters will be considered that form a matrix unlike for surface methods these two parameters are 1 the basic deposit geometry as for surface methods and 2 the support requirement necessary to mine stable stopes or to produce caving a ground control problem boshkov and wright 1973 hamrin 1980 hartman 1987 lewis and clark 1964 thomas 1973 deposit geometry deposit geometry employs the same cutoff points for tabular deposits as in the surface classification but for different

reasons flat deposits require machine handling of the bulk solid at or near the face steep ones can exploit gravity table 6 2 2 with an intermediate inclination recognized if stopes are developed on strike in steep seams as large tunnel sections or step rooms hamrin 1980 machine handling can still be used the resulting stepped configuration causes either dilution or decreased recovery or both because this face can also be benched stope mining simply reproduces tunneling ground control ground control requires knowledge of the structure opening material rock and loads pressures structural components are detailed in table 6 2 10 earlier tables detailed the deposit by its depth and detailed rocks by strength tables 6 2 5 and 6 2 7 respectively from the point of view of support the roof pillars and fill are of primary concern main roof the main roof sometimes the hanging wall is distinguished from the immediate roof by being the critical load transferring element between the overburden and pillars the immediate roof can be removed mined out or supported artificially and lightly the main roof is defined as the first close in competent strong seam if it is only marginally competent heavy artificial support may keep it stable if not then caving can be expected for a flat seam the vertical perpendicular loads on the main roof are largely due to the overburden and its own body load horizontal tangential loads or pressures will tend to be uniformly distributed resulting in a low stress concentration if bed separation occurs above the main roof this stress uniformity is enhanced but at depth overburden loading tends to decrease separation body loads are invariant whereas edge loads particularly those due to the overburden can be shifted pressure arching the main roof is often sufficiently thick so that it can be arched below 1 5 i e at less than 1 horizontally and 5 vertically to increase stability a guideline for coal is that stable spans are usually less than 3 m 10 ft whereas for hard rock they are generally less than 30 m 98 ft for an inclined seam the main roof is the hanging wall and the results are similar to a flat seam pressures perpendicular to it are more significant then tangential ones and bed separation due to gravity is less likely pillars pillars serve to support the main roof and its loads primarily the overburden acting over a tributary area pillar material consists mainly of the seam itself and sometimes waste incorporated within the seam pillars must not only be sufficiently strong but also must be sufficiently stiff a frequently overlooked requirement if pillars are not adequately stiff but still adequately strong the roof will collapse about the still freestanding pillars especially when differential pillar and floor deflection occurs the minimum slenderness ratio for pillars to avoid this crippling is inversely proportional to the recovery th

e mining of flat thick seams of coal dramatically reflects this relationship and is a factor in classifying seam thicknesses table 6 2 4 for massive deposits even in strong rock this makes freestanding pillars of doubtful value upper slenderness ratios range from about 10 1 for coal to 1 3 for rock continuous vertical pillars are used to separate vertical stopes in hard rock that employ steep tabular stoping methods even with stable ground these are usually filled soon after mining for long term stability when massive deposits along with their cap rock are weak caving is necessitated usually performed as horizontal lifts or as block caving caving always requires a sufficient span 9 m 30 ft good draw control and also risks dilution and or poor recovery soft or nonuniform floors footwalls act the same as do soft and irregular pillars fill fill often a sandy slurry consisting of crushed waste cement and water can be readily introduced into confined plugged inclined and steep tabular stopes when drained and dried this hardened slurry provides permanent resistance to ground movement especially for the walls or pillars it is widely used in all but the caving methods it is either run in progressively as a stope is mined out or done all at once at the end of stope mining because of settlement and shrinkage away from a flat back it is marginally useful for flat deposits when timbering is densely placed especially with square sets it rivals pillars it too is usually filled as stoping progresses overhand mining these relationships are summarized in table 6 2 11 and lead into the formal classification underground mining classification system based on an understanding of bulk handling and ground control the underground classification system shown in table 6 2 12 closely follows previous ones the primary difference is that sometimes shrinkage stoping is considered self supported rather than supported however although the broken mineral provides a working floor it is still supporting the hanging wall roof on the other hand when the stope is drawn empty it remains substantially self supported until fill is introduced the disadvantages of the shrinkage method are unique 1 an uncertain working floor 2 dilution due to sloughing and falls of rock 3 possibly adverse chemical effects and 4 tying up about two thirds of the mineral until the stope is drawn vertical crater retreat mining is included in the classification between sublevel and shrinkage stoping hamrin 1980 other factors while subordinated there are additional factors that must be closely evaluated these deal with the broad impacts on the environment health and safety costs output rate and others they are usually evaluated on a relative basis although numbers may also be employed table 6 2 13 boshkov and wright 1973 hartman 1987 an example of where the environmental considerations on the surface are beginning to affect

mining methods is in the use of high density paste backfilling in order to return most of the tailings back underground in order to obtain mining permits from environmental agencies underground in order to obtain mining permits from environmental agencies improvements in geological knowledge for a coal seam within sedimentary rocks that have not been materially disturbed by faulting are unlikely to cause the mining method selection process to be revisited whereas additional geological data drill hole sampling of a steeply dipping tabular gold deposit substantially affected by faulting folding or shearing and displaying pronounced grade trends may require a wider range of possible mining methods to be considered the method of preparing the geological model will also be a significant consideration for the mine planner most of the general mine planning software packages currently available provide powerful three dimensional 3 d visualization wireframe triangulation facilities and block modeling systems to facilitate the development of sophisticated and often complex geological models these tools commonly have the facility for preparing long sections illustrating grade isopachs thickness isopachs and structurally controlled grade trends which can be useful in mining method selection processes interpolation of grade distributions applying simple polygonal methods through to complex geostatistical methods incorporating uncertainty are now common mineral resource models that reflect the inherent uncertainties provide enhanced assistance for the optimum selection of a mining method strength and character of the rock mass this section has been adapted from hoek 2007 any process intended to aid the selection of an excavation method must consider the strength and character of the host rock mass one of the more complex tasks for the mine planner is the determination of representative mechanical properties of the host rock mass although tests have been devised to quantify many of the properties of laboratory rock specimens it is a considerably more difficult task to predict the expected behavior of a rock mass numerous empirical rock mass classification methods derived from actual case studies have been devised to assist mine planners it is important to understand the limitations of rock mass classification schemes palmstrom and broch 2006 and that their use does not and cannot replace some of the more elaborate design procedures or decisions made from economic analyses however the use of these design procedures requires access to relatively detailed information on in situ stresses rock mass properties and planned excavation sequence none of which may be available at an early stage in the project as this information becomes available the rock mass classification scheme adopted should be updated and used in conjunction with site specific analyses open pit slopes the stability of rock slopes has traditionally been e

valuated by limiting equilibrium methods hoek and bray 1981 wyllie et al 2004 although probabilistic based approaches are increasingly more commonly applied because they acknowledge the implicit uncertainties of limit equilibrium methods limit equilibrium models fall into two main categories 1 models that deal with structurally controlled planar or wedge slides and 2 those that deal with circular or nearcircular failure surfaces in homogenous materials many of these models have been available for more than 25 years and can be considered reliable slope design tools wyllie et al provide a methodology for assembling basic geological data rock strength information and groundwater observations and integrating this with engineering rules in the form of design charts and graphical methods to permit a nonspecialist engineer to obtain approximate answers suitable for assessing open pit alternatives several rock mass classification systems have been specifically adapted for rock slope engineering haines and terbrugge 1991 romana 1995 chen 1995 these methodologies have been adapted from classification systems for the highly confined rock mass conditions associated with underground mining as distinct from the low confining stress conditions characteristic of open pit slopes these systems if used with appropriate caution are useful in specifying a range of slope conditions to assist in mining method selection practices but can never replace the requirement for more rigorous processes such as limit equilibrium and numerical modeling of slopes figure 6 3 1 numerical modeling of slope deformation behavior is now a routine activity on many large open pit mines software programs such as flac and udec are typically used for such modeling although a significant amount of expertise is required to ensure realistic input information and reliable interpretation of outputs in best practice a combination of limit equilibrium and numerical modeling approaches are applied to generate an array of solutions for the range of inputs that typically exist at a site because it is far more reliable to look at the array of results from a parametric study than a single deterministic study with the greater depths characteristic of modern openpit mines the role of the in situ stress field in slope stability is becoming an increasingly important consideration in these cases mine planners must seek advice from specialists about the applied assumptions when comparing deep open pit alternatives with underground methods in terms of arriving at a suitable set of slope parameters for assessing the applicability of any open pit method a process that recognizes the implicit uncertainties and considers a range of slopes as inputs to the evaluation should always be adopted underground excavations rock mass classification systems applicable to underground excavations have been evolving for more than 100 years since ritter 1879 attempted to formalize an e

mpirical approach to tunnel design for the purposes of determining support requirements terzaghi s rock mass classification the earliest reference to the use of rock mass classification for the design of tunnel support is in a paper by terzaghi 1946 in which the rock loads carried by steel sets are estimated on the basis of a descriptive classification it is useful to examine the rock mass descriptions included in his original paper because he draws attention to those characteristics that dominate rock mass behavior particularly in situations where gravity constitutes the dominant driving force the clear and concise definitions and the practical comments included in these descriptions are good examples of the type of engineering geology information that is most useful for engineering design terzaghi s descriptions quoted directly from his paper are as follows intact rock contains neither joints nor hair cracks consequently if it breaks it breaks across sound rock because of injury to the rock due to blasting spalls may drop off the roof several hours or days after blasting known as a spalling condition hard intact rock may also be encountered in the popping condition involving the spontaneous and violent detachment of rock slabs from the sides or roof stratified rock consists of individual stratum with little or no resistance against separation along the boundaries between the strata the strata may or may not be weakened by transverse joints in such rock the spalling condition is quite common moderately jointed rock contains joints and hair cracks but the blocks between joints are locally grown together or so intimately interlocked that vertical walls do not require lateral support in rocks of this type both spalling and popping conditions may be encountered blocky and seamy rock consists of chemically intact or almost intact rock fragments which are entirely separated from each other and imperfectly interlocked in such rock vertical walls may require lateral support crushed but chemically intact rock has the character of crusher run if most or all of the fragments are as small as fine sand grains and no re cementation has taken place crushed rock below the water table exhibits the properties of a water bearing sand squeezing rock slowly advances into the tunnel without perceptible volume increase a prerequisite for squeeze is a high percentage of microscopic and submicroscopic particles of micaceous minerals or clay minerals with a low swelling capacity swelling rock advances into the tunnel chiefly because of expansion the capacity to swell seems to be limited to those rocks that contain clay minerals such as montmorillonite with a high swelling capacity lauffer 1958 proposed that the stand up time for an unsupported span is related to the quality of the rock mass in which the span is excavated in a tunnel the unsupported span is defined as the span of the tunnel or the distance between

the face and the nearest support if greater than the tunnel span lauffer s original classification has since been modified by a number of authors notably pacher et al 1974 and now forms part of the general tunneling approach known as the new austrian tunnelling method this method includes a number of techniques for safe tunneling in rock conditions in which the stand up time is limited before failure occurs these techniques include the use of smaller headings and benching or the use of multiple drifts to form a reinforced ring inside which the bulk of the tunnel can be excavated these techniques are applicable in soft rocks such as shales and phyllites and in which the squeezing and swelling problems described by terzaghi are likely to occur the techniques are also applicable when tunneling in excessively broken rock but great care should be taken in attempting to apply these techniques to excavations in hard rocks in which different failure mechanisms occur rock quality designation rqd the rqd index was developed by deere et al 1967 to provide a quantitative estimate of rock mass quality from drill core logs rqd is defined as the percentage of intact core pieces longer than 100 mm in the total length of core the core should be at a minimum size 54 7 mm in diameter and should be drilled with a double tube core barrel the procedures for measurement of the length of core pieces and the calculation of rqd are illustrated in figure 6 3 2 rqd is a directionally dependent parameter and its value may change significantly depending upon the borehole orientation the use of the volumetric joint count can be quite useful in reducing this directional dependence rqd is also intended to represent the rock mass quality in situ when using diamond drill core care must be taken to ensure that fractures which have been caused by handling or the drilling process are identified and ignored when determining the value of rqd deere s rqd was widely used particularly in north america after its introduction although various investigators have sought to relate rqd to terzaghi s rock load factors and to rock bolt requirements in tunnels the most important use of rqd is as a component of the rock mass rating and q rock mass classifications covered later in this chapter rock structure rating rsr wickham and tiedemann 1974 described a quantitative method for describing the quality of a rock mass and for selecting appropriate support on the basis of their rsr classification most of the case histories used in the development of this system were for relatively small tunnels supported by means of steel sets although historically this system was the first to make reference to shotcrete support although the rsr classification system is not widely used today wickham et al s work in which they devised a basis for rating the geological geometrical and joints condition played a significant role in the development of the classifica

tion schemes discussed in the remaining sections of this chapter rock mass rating rmr bieniawski 1973 1976 published the details of a rock mass classification called the geomechanics classification or the rmr system over the years this system has been successively refined as more case records have been examined and the reader should be aware that bieniawski has made significant changes in the ratings assigned to different parameters bieniawski 1989 bieniawski s rmr system was originally based on case histories drawn from civil engineering consequently the mining industry tended to regard the classification as somewhat conservative and several modifications have been proposed in order to make the classification more relevant to mining applications a comprehensive summary of these modifications was compiled by bieniawski 1989 both this and the 1976 version deal with estimating the strength of rock masses the following six parameters are used to classify a rock mass using the rmr system 1 uniaxial compressive strength of rock material 2 rqd 3 spacing of discontinuities 4 condition of discontinuities 5 groundwater conditions 6 orientation of discontinuities in applying this classification system the rock mass is divided into a number of structural regions and each region is classified separately the boundaries of the structural regions usually coincide with a major structural feature such as a fault or with a change in rock type in some cases significant changes in discontinuity spacing or characteristics within the same rock type may necessitate the division of the rock mass into a number of small structural regions modified rock mass rating mrmr laubscher 1977 1984 laubscher and taylor 1976 and laubscher and page 1990 have described an mrmr system for mining this system takes the basic rmr value as defined by bieniawski and adjusts it to account for in situ and induced stresses stress changes and the effects of blasting and weathering a set of support recommendations is associated with the resulting mrmr value in using laubscher s mrmr system it should be borne in mind that many of the case histories upon which it is based are derived from caving operations originally block caving in asbestos mines in africa formed the basis for the modifications but subsequently other case histories from around the world have been added to the database the selection of an appropriate mass underground mining method has been presented by laubscher 1981 figure 6 3 3 the selection process is based on his rock mass classification system which adjusts for expected mining effects on the rock mass strength laubscher s scheme is aimed at the mass mining methods primarily block caving and open stoping methods although his main emphasis is on cavability the two parameters that determine whether a caving system is used over a stoping system are the degree of fracturing rqd figure 6 3 3 joint spacing an

d the joint rating which is a description of the character of the joint that is waviness filling and water conditions this scheme puts emphasis on the jointing as the only control for determining cavability laubscher 1990 has subsequently modified the mrmr classification to relate the mrmr rating to the hydraulic radius hr figure 6 3 4 by including the hydraulic radius cavability becomes feasible for more competent rock if the area available for undercutting is large cummings et al 1982 and kendorski et al 1983 have also modified bieniawski s rmr classification to produce the modified basic rmr mbrmr system for mining developed for block caving operations in the united states this system involves the use of different ratings for the original parameters used to determine the value of rmr and the subsequent adjustment of the resulting mbrmr value to allow for blast damage induced stresses structural features distance from the cave front and size of the caving block support recommendations are presented for isolated or development drifts as well as for the final support of intersections and drifts rock tunneling quality index q on the basis of an evaluation of a large number of case histories of underground excavations barton et al 1974 of the norwegian geotechnical institute proposed a rock tunneling quality index q for the determination of rock mass characteristics and tunnel support requirements modified rock quality index q barton s q has been used with a great deal of success in the design of tunnels in rock however the srf parameter is redundant when the classification system is used for the estimation of rock mass properties for the purpose of analytical or numerical modeling for design because the influence of stress is taken into account within the model thus the srf is set to 1 0 which is equivalent to a moderately clamped but not overstressed rock mass in addition in most underground hard rock environments the excavations are relatively dry not considering transient mine water inflows from drilling or backfilling in which case the jw parameter can also be set to 1 0 along with several other factors accounting for jointing stope geometry and mining induced stress q can then be used to determine the modified stability n which is then used with the modified stability graph method mathews et al 1981 potvin 1988 bawden 1993 and hoek et al 1995 for the dimensioning of open stopes in underground mines to facilitate the following discussion on mining method selection and evaluation methodologies a brief discussion to characterize the range of surface i e open pit and underground mining methods follows additional information in relation to open pit mining methods can be found in hustrulid and kuchta 2006 and kennedy 1990 while further information on underground mining methods can be found in hustrulid and bullock 2001 surface mining methods surface mining methods

are defined here as any excavation that commences from the natural surface and does not entail the construction of a tunnel or shaft most often the style of mineralization will significantly impact the features of a surface mining method particularly the character and thickness of overburden the type of equipment deployed also commonly affects the classification of a surface mining method open cut mining open cut mining refers to a particular kind of surface mining that most commonly deploys large rope shovels hydraulic shovels or excavators together with suitably sized rear dump trucks and progresses the excavation in a series of slices for hard rock open cut mines drilling and blasting practices are often an integral part of the excavation system the mining slice height may or may not be consistent with the vertical interval applied to construct berms on a slope for this method pit slopes commonly emerge as a sequential series of cutbacks designed to manage the progressive strip ratios and maximize cash flows many of these features are illustrated in figure 6 3 6 the term open cut mining is also commonly used to describe excavation in soft rock in which drilling and blasting systems may not be required and continuous excavation technologies such as bucket wheel excavators are utilized truck and shovel systems may well still have a role in these environments strip mining strip mining describes a particular type of surface mining method that relies heavily on the progressive and sequential disposal of overburden spoil into a previously mined void figure 6 3 7 coal mining often falls in this category although bauxite miners often adopt a variant of strip mining dragline equipment supplemented by truck and shovel systems are often observed in strip mines strip ratios can be relatively high and slope angles can be relatively steep largely due to the relatively low overall height of these slopes underground mining methods underground mining methods invariably rely on tunneling networks to gain access to the zones of valuable minerals figure 6 3 8 these tunnel networks can be linked to vertical shafts equipped with rock hoisting facilities or inclined ramps also known as declines through which rubber tired equipment can pass both to facilitate movement of this equipment and also for the transport of rock products to the surface underground mining methods can be divided into three broad classes caving stoping and other methods the term caving implies the controlled collapse of the rock mass under the force of gravity whereas the term stoping implies the excavation of a stable opening of small or large dimensions caving methods three generic methods of removing the valuable minerals and triggering caving processes can be described 1 block caving technologies are suitable for large lowgrade ore bodies either vertical or inclined which are undercut over a large area thereby inducing collapse of the entire ro

ck mass with the broken rock being extracted via a purpose constructed system of extraction points figure 6 3 9 the collapsing rock mass usually propagates to the natural surface and requires careful draw management to contain dilution from unmineralized material large scales of operation are possible with this technology which is successfully being applied at increasing depths and on increasingly stronger rocks compared to reference points in the 20th century 2 sublevel caving technologies are also suitable for large ore bodies of a generally tabular geometry requiring a more selective mining system and extraction by conventional drilling and blasting technologies figure 6 3 10 this method differs from block caving in that all of the ore is drilled and blasted and only the overburden waste rock caves by gravity depending on the ore body geometry this mining method is amenable to high rates of ore production sublevel caving is extensively used in the swedish iron ore mines at kiruna 26 mt a 3 longwall mining systems are extensively applied to deposits of coal and rely on rock cutting technologies to excavate the coal using shearers and conveyor transportation systems to deliver the coal to surface coal seams are removed in a single slice which may be 300 to 400 m across and several kilometers long with working heights of between 1 5 and 6 m the overlying rock collapses into the mining void as the shearing system is advanced the shearing system is protected by a series of heavy duty shields operated by means of hydraulic jacks which provide a movable canopy thereby preventing the immediate roof over the workings from collapsing onto the shearer recent innovations also include longwall top coal caving which emphasizes coal extraction efficiency in thicker seams stoping methods the excavation of a stable void of small or large dimensions may be stoping s first defining feature that is the shape of the mineralization and or the nature of the rock mass in which the excavation is to be constructed the term stoping infers the excavation of a stable void of small or large dimensions and is a defining feature of these methods table 6 3 1 the character of this void is substantially influenced by the shape of the mineralization and or the nature of the rock mass in which the stope is being constructed a second defining feature of the range of possible stoping methods comprises the use of a fill material which typically falls into two main groups 1 waste rock or tailings which may either be unconsolidated and therefore have negligible strength or 2 consolidated typically with a pozzolanic material such as cement or fly ash in this context the term strength implies a capacity of the fill material to stand without collapse when otherwise virgin rock confining the fill material to the original void is removed in a second phase of mining a third defining feature of stoping methods is the nature of the drilling and bl

asting technology deployed which may be characterized as either short hole or long hole systems in this case short holes are typically less than 4 m in length and consistent with a single pass tunneling jumbo while long holes are drilled with purpose designed long hole drilling machines applying segmented drill strings a fourth defining feature occurs where numerous independent stopes and the sequence of creation and possibly filling of these voids are contemplated collectively these features give rise to a large number of stoping methods as in table 6 3 1 including shrink stopes figure 6 3 11 and overhand cut and fill stoping figure 6 3 12 a fifth defining feature is the degree to which mineralization that has economic value is not mined so as to maintain the required stability of the stope thus the principle of pillars such as in the room and pillar method commonly applied to mineralization with low dips and modest heights 4 6 m see figure 6 4 1 in chapter 6 4 however it can be applied to heights of more than 30 m vertical crater retreat vcr is a term that encapsulates open stopes being developed by applying a particular drilland blast methodology but is otherwise similar to conventional open stoping figure 6 3 13 other methods the extraordinary diversity of mineral deposits inevitably leads to innovative mining methods including combining methods one such method is postpillar cut and fill where the pillars are mined small but immediately backfilled another example is the avoca method which combines sublevel longhole drilling with immediate backfilling other approaches are described in the following paragraphs horodiam the horodiam method utilizes a large diameter raise bored shaft as the access for a drill jumbo to drill horizontal radial blastholes over the entire height of the stope the method is amenable to remote control technologies and has been patented as a remotely operated excavation system roes by australia s commonwealth scientific and industrial research organisation coal seam methane drainage the success of this method is a function of the physical properties of the coal seam diffusivity reservoir pressure permeability and gas content mining method if in progress and drainage method hartman 1987 the method is used in europe but only now is gaining widespread acceptance in the united states to produce coal seam methane this method is a type of borehole mining in which the wells are used to recover the methane coal seam gasification this method is applied to coal and is also related to borehole mining in which the coal is burned at one end and the gases given off are recovered at another borehole use of this method is based on whether the cost of burning the coal and recovering the gases is cheaper than traditional mining the key parameters that impact the method are the fracturing and the chemical composition of the coal this method may become more feasible in cases wh

ere the coal seam is too narrow for traditional methods in the recovery of multiple seams where the second seam is too close to the first to be recovered in a traditional fashion or where seam depth or quality precludes the economic application of conventional surface or underground methods subsidence considerations apply as for caving methods of mining underground retorting this method is being tried with oil shales and tar sands after the area is mined to some extent using traditional drifting techniques and pillar designs the rock in the stope retort is blasted in place oil is released from the rock and recovered under the stope this method is chosen based on the retorting characteristics rather than on the mining parameters from a mining perspective the critical factor would be the cost and methodology of fracturing the ground the degree of fragmentation will impact the percentage recovery of the oil which is probably the most critical concern surface to underground transition methods occasionally newly discovered mineral deposits are amenable to both a surface mining and an underground method which presents a particularly interesting challenge to mine planners typically a surface mining method would be applied to initially develop the deposit although instances of an underground development preceding an open pit development have occurred the answer to critical questions regarding the depth at which a transition should occur is dependent on many factors including the relative scale of the surface mine and the underground mine the lead time required to develop the underground mine and the optimum underground mining method mining method selection and evaluation methodologies in many cases the style and geometry of the mineralized system will be the dominant factor in identifying the most appropriate mining method for evaluating the potential economic value of the deposit it is uncommon however to encounter a mineral deposit that is amenable only to a single mining method consider for example a flat tabular potash deposit within a halite sequence occurring at a depth that unequivocally precludes open pit methods in this instance solution mining and conventional room and pillar mining systems should both be considered during the initial appraisal and carried forward in subsequent appraisals until such time as a clear economic benefit from a preferred method after accounting for risk and uncertainty can be demonstrated a second factor that will always have substantial influence on the possible mining methods is the characteristics of the host rock mass and the mineralized rock mass the quantity and quality of this information will almost certainly be influenced by the status of exploration over the deposit an all too common problem is the failure in early stage exploration activities to allocate sufficient funds for the collection of critical rock mechanics data by which the characteristics of the rock ma

ss can be adequately ascertained the techniques for evaluating mining methods are only attempts at defining and quantifying in a written format what miners in years past determined through discussion previous experience and intuition therefore each of the method selection schemes presented here is similar and yet different reflecting personal preferences in their emphases the purpose of discussing these techniques is not to critique them but simply to present the alternatives available to aid in selecting the most appropriate most of the schemes are aimed at determining the appropriate underground method as there are many possible choices however the purpose of this chapter is to discuss the selection of the best mining methods including surface hydraulic and more novel methods the method selection process should first determine whether the deposit should be mined using a more traditional surface underground or in situ leach mining method a novel method should only be considered if traditional methods are not economically or technically feasible to start a mine with a novel mining method requires adequate funding and an enormous commitment from the board of directors to technical development the board must also have the patience to work out the technical problems if the deposit cannot be mined using a surface method then an underground method should be considered the mining method selection techniques are limited because selection is based solely on the known physical parameters and rock strength characteristics sometimes several mining methods may appear to be equally feasible in order to further determine which method s is the most suitable the input variables of mining costs mining rate labor availability and environmental regulations should be considered in more detail note none of the mtethod selection systems deal with in situ stress although the techniques account for the vertical stress via depth none of the methods discuss how a high horizontal stress impacts the choice of the mining method risk increasingly business investment decision making requires quantification of the risks and uncertainties associated with a mine plan and particularly with the value of the proposed investment in which an investment opportunity is competing against many other investment opportunities ample evidence within the global mining industry demonstrates that the highestranked risk to investment value is the quantity and quality of the mineral resource the key revenue drivers on which deterministic mine plans are based consequently the trend is toward the preparation of probabilistic resource models in mine planning evaluations both to describe the probability weighted estimate of the value of a mineral deposit constrained by deterministic limits and to describe the value of a mineral deposit reflecting the probabilities of different mineral limits and grade distributions these latter techniques sometimes kno

wn as resource range analyses are particularly useful to investment decision making processes in the early stages of exploration of a mineral deposit and are supplanted by the probabilistic resource models as the exploration process matures in its simplest form a resource range analysis for a mineral deposit comprises the development of five equally plausible mineral resource models sometimes characterized as the minimum low most likely high and maximum cases with probabilities assigned to each case such that the cumulative probability equals unity in some instances three models reflecting the low most likely and high cases may suffice similarly a mine plan for each model of the mineral resource is prepared and evaluated with a probability weighting to derive an expected value the different resource models may well dictate that materially different mining methods are selected for the evaluation the significant increase in mine planning effort arising from these approaches manifestly relies on the emergence of numerical evaluation methodologies previously described economic analysis every corporation has a range of metrics in which it is interested when evaluating a mineral deposit these could include issues such as scale and the level of participation in the market for the product the likely life of the mining project the size of the investment required to bring the deposit into production and similar physical metrics invariably this must also include financial metrics with little exception the value of prospective future cash flows discounted at an appropriate rate as a present value and netted against the investment required so as to describe a net present value npv will also be required particularly where competing investment alternatives must be assessed in conducting any economic analysis of a mineral project definition of the future cash flows should consider the four basic cash streams 1 product income the incoming cash stream that describes income from the sale of product perhaps net of royalties as may be considered appropriate 2 capital cost expenses the cash stream representing the outgoing expenditures required to develop the mining project and create an income stream sustaining capital expenditures should also be regarded 3 operating cost expenses the outgoing cash stream representing the expenditure on inputs to the production process incurred to produce units of output and generate an income stream 4 tax expenses the outgoing cash stream representing the payment of income taxes and any offsetting depreciation allowances an accumulation of these four cash streams in any accounting period typically years provides the basis for describing discounted cash flows and npv probabilistic evaluations should also be completed applying stochastic simulations for the income stream and mine plans that examine optimistic most likely and pessimistic scenarios for the capital cost and operating

cost streams stages in application of mining method selection process application of any mining method selection process is necessarily iterative and as knowledge about the deposit under consideration evolves may have different conclusions deposit discovery the first occasion on which evaluation practitioners may be called upon to provide a business case to justify further exploration expenditure is shortly after the emergence of promising exploration results enthusiasm for further drilling is inevitably high and depending on the prevailing economic environment and the quality of the exploration results may overshadow a more rational assessment of the potential economic value this is an ideal time to apply a resource range analysis methodology to establish a range of resource models and conceptual mine plans for each model and more importantly to establish the business case for further investment in exploration relative to the corporation s business goals and required deposit size in this circumstance the mineral resource models are likely to be simple polygonal models that give some sense of the shape and geometry the conceptual mine plans are unlikely to be any more than high level estimates of output that have been extended by relevant unit costs for the expected activities perhaps applying many rules of thumb for mining and processing activities advanced exploration concept studies the term advanced exploration here implies that preliminary mineral resource modeling has been completed and that the mineral resource estimates emerging from the process are able to be classified by one of the internationally recognized systems typically at least as an inferred mineral resource the presumption is that the mineral resource model is based at least on block modeling technology with interpolated mineral concentrations with the model extended to describe the surrounding host rock for the purpose of modeling slopes in the case of deposits requiring open pit evaluations in this case one of the readily available open pit optimization software systems should be applied to evaluate the deposit for underground deposits similar systems are starting to become available and these will evolve over time with industry research and development programs many unvalidated critical assumptions may well need to be made about a number of mining method alternatives derived by applying the methodologies previously outlined because of the relatively high levels of uncertainty surrounding many of the key drivers resource range analysis methodologies are extremely useful in helping to establish the impacts of uncertainty on project value and identifying the priorities for resolving these uncertainties to more acceptable levels in most circumstances a risk weighted npv greater than zero should be sufficient to encourage further investigation by way of preliminary feasibility studies preliminary feasibility studies in a mineral project developm

ent the advance to a preliminary feasibility study usually signifies a stronger commitment to drilling and sampling the mineralization complemented by further rounds of mineral resource modeling hydrogeological geotechnical geometallurgical and environmental studies are also likely to proceed welded together under a strategic mine planning umbrella improved resource estimate confidence is likely mining method studies are likely to be of less significance as the focus is on variations within the selection of methods identified at the concept study stage in most cases the study effort is now focused on ensuring that all reasonable alternatives are investigated with the intent of identifying the best alternative to move forward to a final feasibility study final feasibility studies for final feasibility studies the focus usually shifts from mineral resource definition to detailed mine and infrastructure planning at a level that is adequate for the capital expenditure approvals necessary to take the project into construction evaluation of alternative mining methods should never figure in this stage of project evolution the soft rocks usually are part of the sedimentary minerals classification which is subdivided into clastic organic and chemical examples of the soft rock ores include coal metalliferous shales oil shales potash salt trona and possibly kimberlites where coal metalliferous shales potash and trona occur as economic ores they are typically laterally extensive beds in a nearly horizontal inclination but with at most a shallow dip angle this differentiation is key because it enables the application of large scale mechanization to the mining process the economy of scale that results from mechanization is often the determinant factor for economic success as the capabilities of mechanical cutting expand into more demanding applications the possibility exists that ores previously considered hard rock deposits such as the platinumbearing reefs in southern africa may be cut instead of blasted coal has its origins in the accumulation of plant debris that becomes buried by sediments through a process dependent on time burial depth and chemical transformation the plant debris becomes coal therefore coal is classified as an organic sedimentary mineral coal varies in quality from lignite to anthracite with sub bituminous and bituminous ranked as intermediate in the progression and the most commonly mined types worldwide coal is primarily used for electricity generation and steelmaking and is commonly referred to respectively as steam coal and coking coal coal with attributes such as an appreciable free swelling index fsi is used to make coke which is used in primary production of steel from iron ore the scarcity of such coals elevates their value compared to steam coal trona is a carbonate mineral of sodium used to form soda ash used in glassmaking and other industrial processes including bak

ing soda it occurs naturally in a few locations worldwide as laterally extensive evaporite deposits suitable for underground mining methods it is a moderate value ore and competes with an alternative process that synthesizes the same product from chemical feedstocks potash is the name loosely applied to a variety of potassium salts particularly potassium chloride which are encountered in laterally extensive evaporite deposits found worldwide these are often associated with intermixed or stratigraphically adjacent halite sodium chloride potassium is key to plant growth and potash is mainly used as fertilizer although it is also used to produce soaps ceramics and drugs among others coal potash trona and salt are the principal soft rock ores and within limits share similar production methods that focus on economies of scale the most common mining techniques for soft rock ores are longwall room and pillar r p and stope and pillar for water soluble minerals solution mining is an alternative the process of properly selecting an underground mining method for a particular ore deposit is critical to the ultimate success of the operation an improperly selected method will increase costs lower productivity create unnecessary hazards and reduce resource recovery due to the complex nature of ore bodies no two mines are completely alike and all operations must adapt to the particular conditions of their deposits ore deposit characteristics numerous considerations must to be recognized when selecting the best method to mine a soft rock ore deposit some of the considerations are based on ore deposit characteristics favorable to the mining method being considered ore strength host rock strength deposit shape deposit dip deposit size deposit thickness deposit grade ore uniformity deposit depth other characteristics are a function of mining method operating cost capital cost and development timing production rate mechanization selectivity and flexibility health and safety environmental effects ore strength the material properties of the ore often drive mine design decisions although there are many mechanical properties compressive strength is often discussed as an indicative characteristic inferring structural performance and suitability for mechanical cutting mining methods such as r p and stope and pillar depend on the strength of the ore rock to support the roof and overburden in order to create a structurally stable excavation in soft rock applications the relative strength of the ore is often weak with a compressive strength less than 6 000 psi this low strength is generally associated with a low to moderate specific energy of cutting kilowatt hour ton this allows the application of mechanized cutting and loading which is elemental to the success of many modern mines as the ore strength increases the options for mechanical cutting are reduced and the application productivity

declines while costs increase there is a marked difference between the cost and productivity performance of mechanized cutting versus drilling and blasting in the majority of soft rock applications with mechanized methods decidedly preferred table 6 4 1 gives the strength designations and ranges of values based on the compressive strength of the material it is important to note that the strength and mechanical properties of a rock are significantly affected by fracturing and planes of weakness in the deposit fracturing is characterized by small discontinuities in the rock mass and may be caused by heat vapor expansion as in porphyry deposits depositional conditions i e slickensides or tectonic movement faults cleat is a fracture system ordinarily observed in coal two different fracture directions are typically present face cleat primary direction and butt cleat secondary direction during exploration the degree of fracturing should be quantified and utilized to reduce ore structural properties potentially leading to smaller openings larger pillars and increased ground control costs limited fracturing may be a positive factor for some mining methods because it promotes caving lowers blasting requirements and aids mechanical cutting however excessive fracturing can have a negative influence on ground control water and gas inflows host rock strength the strength of the rock enclosing the ore is also an important driver in mining method selection temporary and permanent openings must be developed either in the host rock in order to access the ore or with the host rock as roof back or hanging wall or floor footwall for the ore openings entries or crosscuts to execute an appropriate design the material properties must be understood the behavior of the roof and floor can be pivotal in the success of mechanized mining systems floors that become muddy and easily rutted can disable production and send maintenance costs skyrocketing it is inaccurate to assume that the ore and host material will have the same characteristics so each must be independently characterized by geomechanical testing deposit shape ore deposits are classified into two broad categories tabular and massive a tabular deposit is flat and thin and has a broad horizontal extent this classification typically refers to materials formed by sedimentation similar in shape to tabular ore bodies lenticular deposits are shaped like lenses and lack the large areal extent of most tabular deposits most methods designed to exploit tabular deposits may be adapted to mine lenticular ones the ore materials must often be of higher value than applications such as coal because production costs are generally higher but reserve tonnages are lower a massive deposit may possess any shape the ore is often distributed in low concentrations over a wide area with varying horizontal and vertical extents frequently the difference between ore and waste m

ay be a function of grade rather than rock type massive deposits may be unpredictable and require a considerable exploration investment in order to document and fully understand the resource for the purposes of mining method selection massive deposits are often accompanied by a more specific clause like massive with large vertical extent these additions are necessary because the shape of a massive deposit is variable and may be unsuitable for certain mining methods the deposit shape definitions are summarized in table 6 4 2 deposit dip dip is defined as the angle of inclination of a plane measured downward perpendicular to the strike direction the deposit dip is more relevant to tabular ore bodies than massive ones although it may sometimes be a consideration for the latter deposit dips are categorized and defined in table 6 4 3 both flat lying soft rock ore beds and near vertical ore veins may be classified as tabular but the mining methods used to exploit them are dramatically different several methods are highly dependent on gravity for material flow and cannot function in flat lying deposits alternatively low working slopes are a key factor in the application of mechanization for cutting and loading as well as material haulage by rubbertired rail or conveyor belt methods deposit size the volumetric size of an ore body must also be considered several of the methods discussed in this chapter rely on large deposits with long mine lives to justify their high initial capital costs and promote economies of scale other methods simply do not work efficiently in ore bodies which are either too large or too small deposit size is characterized subjectively by the terms small medium and large as a generalization large ore deposits have tens to hundreds of million cubic yards of ore and suggest mine lives in the 10 to 50 year range deposit thickness deposit thickness refers to the ore thickness of tabular deposits thickness plays an important role in opening stability and may prevent certain equipment from functioning efficiently or mining methods from being effective the deposit thickness nominally the mining extraction height definitions are listed in table 6 4 4 these definitions are most relevant to mechanical cutting and loading applications such as longwall or continuous miner r p the thickness ranges roughly correlate with the types of equipment available to implement a mining system and the cost productivity that might be expected deposit grade grade is discussed in terms of the amount value of recoverable salable material in a unit weight or volume of in place mineral resource where it becomes economically viable to produce the mineral resource the in place resource becomes ore as such the end outcome economics of different mining methods may vary the amount of ore that an in place mineral resource may yield a gold ore may contain as low as 0 1 oz ton and still be economic whereas iron ore grades may ap

proach 60 by weight coal is generally characterized by its attributes that is energy content btu lb percentage of ash moisture and sulfur fsi and so forth some mining methods with high operating costs necessitate high grade ores in order to be economic large scale methods may be suitable for large low grade deposits such as bituminous coals ore grades are categorized subjectively and must be investigated on an individual site basis ore grade definitions are provided in table 6 4 5 value estimates associated with the classifications give some relative sense of the range involved ore uniformity the uniformity of the ore in the mineral deposit must be considered as poor uniformity may render some mining methods unviable it is undesirable to excavate subeconomic material unless it is necessary to reach ore or create necessary infrastructure such as belt conveyor galleries a mineral deposit may be segmented by faults subeconomic mineral occurrence or legal environmental issues some mining methods are well suited to flexibility because they can selectively extract specific sections of a deposit without disrupting the overall operation an example of this is the case where an r p coal mine adapts the panel geometry while in panel to reflect new findings about unsatisfactory coal quality adverse roof conditions or insufficient coal thickness other methods such as longwall mining limit selectivity and must produce at least some amount of material leading to equipment advance in order to continue to the panel s intended end faults with significant displacements compared to the bed thickness can seriously disrupt longwall or r p operations in some areas igneous or sedimentary materials may be injected into tabular deposits such as dolerite dykes in coal seams and create impediments to mechanized cutting and loading an inconsistent feed of material may disrupt processing plant performance or require blending rehandling or disposal of mined material these situations can be anticipated and minimized with a thorough knowledge of the ore body s uniformity ore uniformity designations are variable moderate fairly uniform and uniform deposit depth another deposit related consideration that impacts mining method selection is ore deposit depth relative to the surface shallow deposits are generally more suited for surface mining deeper deposits may require progressively greater ground control measures increased costs larger pillar sizes lower recovery or decreased applicability of some mining methods in order to ensure safety and sustainability commonly applied variations of r p or longwall mining occur over deposit depth ranges from 250 to 3 500 ft the definition of shallow moderate deep is relative depending on the value of the ore and the strength of the material a deep coal mine might have workings to a depth of 3 500 4 500 ft alternatively a deep gold mine producing from a meta quartzite reef migh

t have workings to nearly triple that depth classification for deposit depths are shallow moderate and deep mining method characteristics every mining method has characteristics that will produce different outcomes based on the ore deposit to be mined as such prior to selecting the best mining method the methods to be applied and their expected outcomes must be clearly understood operating cost the operating cost of a mine is the cost associated with the production of ore from the primary mining method the total cost is higher and incorporates items such as depreciation depletion taxes and royalties the operating cost divided by the number of salable units of production mined creates a metric used to compare efficiency between competing production alternatives that is ton when the total cost is the basis of the metric it can indicate the potential viability of the project in total in mining the operating cost is composed of fixed and variable expenses variable expense totals change in proportion with activity such as roof control cost ft that typically accumulates with the amount of entry development in comparison fixed costs such as labor and ventilation stay relatively constant over a moderate range of activity variation some methods are labor intensive or may require a large quantity of materials in order to operate thereby necessitating valuable ores to compensate for the greater price of extracting them other methods cost little once implemented but have high initial capital costs these methods such as longwall mining may be able to excavate large low grade deposits economically capital cost and development timing initial capital cost is defined as the amount of investment needed before the mine begins to generate revenue a small quarry excavating an outcropping limestone bed has little capital cost because it can start extracting ore almost immediately with little investment in equipment alternatively a deep potash mine might have to sink one or more shafts beyond a depth of 3 000 ft build a surface plant and implement a mechanized mining equipment fleet to produce the first salable ton of product thus first production may come after several years and tens to hundreds of million dollars have been committed higher capital costs are frequently associated with long development or start up times equipment manufacturers often have wait times of months or even years before assembly and delivery of new equipment typically this equipment is customized for the mine specific application production rate the production rate of a mine is highly dependent on the mining method a high production rate can accommodate a large market and may overcome low value ore if operating costs are low the ability to stockpile and blend ores of varying grade in order to maintain a consistent feed to the mill is typically advantageous higher production is generally more desirable because mines are rarely opened in areas

where selling more product is disadvantageous the economics of mines that can sell product up to the limit of their production capacity are drastically different than mines that can produce at levels above what their markets can consume in the latter case production enhancement proposals readily embraced by the former case intended to distribute fixed costs over a larger total production are rejected and the focus sharpens on costs contributing to the fixed component of operating cost mechanization mechanization is a critical element of a modern mine utilizing machines to perform production tasks is much safer and more efficient in cost or production performance than using manual labor to justify a large capital investment in equipment it is common to need a longer mine life and thus a larger ore body highly mechanized mining is safer than less mechanized methods because fewer workers will be needed and thus the overall hazard exposure will be lower several methods lend themselves to a high degree of mechanization including longwall and continuous miner r p methods selectivity and flexibility selectivity and flexibility can significantly contribute to the success of a mining method it is generally valid to assume that mining conditions market prices and technology will change over the course of a mine s life so the mining method must be adaptable to these fluctuations sacrificing optional alternatives in any mining method is not desirable unless there is compelling reason to do so if commodity prices were to drop substantially a portion of the ore in a massive deposit may become uneconomic to mine if the mining method is able to bypass the low grade sections and continue mining economic material the mine will continue to be successful health and safety the safety and health of a mine s workers should be the top priority of every operator several methods are inherently safer than others because the openings are more stable or personnel are less likely to be subjected to hazardous conditions although no modern methods are considered to be unsafe it bears mentioning that specific health and safety concerns are often mitigated by the mining method selection longwall mining is recognized as the safest method of mining applied to soft rock deposits environmental effects the largest environmental impacts of an underground mine typically fall into three categories subsidence groundwater and atmospheric emissions subsidence is defined as the sinking of the surface above mine workings as a result of material settling into the voids created by mineral extraction it is contentious in urban or suburban areas where it can affect homes schools and roads the surface subsidence created by modern longwall mines is largely predictable in its timing and magnitude in contrast to the unpredictable outcomes associated with some r p mines in this way longwall subsidence is less hazardous to human made surface structures becau

se impacts occur soon after mining and rarely change much after initial stabilization this allows remediation of surface damage in a time contemporary with mining most areas with a history of mining also have developed legal processes to address damage from mining induced subsidence high extraction mining methods will foreseeably induce surface subsidence if selected provisions must exist to mitigate or remedy damages water impacts may arise by accidental causes acidgenerating rock of multiple types in excavated ore waste or overlying strata may produce acid mine drainage water produced by the rock mass and mining process must be afforded appropriate controls as it will be necessary to keep the mine drained in all cases strict controls must be effectively applied to mitigate groundwater or surface water impacts by miningrelated water discharges air quality in underground mines is typically affected by the natural liberation of mine gases i e methane ch4 hydrogen sulfide h2s and carbon dioxide co2 blasting by products and equipment emissions i e nitrogen oxides nox sulfur oxides sox and diesel particulate matter and mineral dust from ventilation fans generally exposure and emission thresholds exist for these emissions and are strictly applied in the case of coal dust and methane special precautions are followed to avoid the hazards of fire and explosion zero harm is a sustainability principle applied by the foremost mining enterprises in the context of health safety environment and communities where mines in their portfolios actively operate it is an acknowledged goal that communities will be forever improved because of the global and local activities of these mining enterprises mine planning mine planning has three well defined stages in order to have a successful implementation of the project and operation of the mine identification selection and definition identification the initial assessment is a review of information about the potential mining site and involves the analysis of geographic geologic environmental technical and economic data this assessment helps the mining company to evaluate the advantages and disadvantages of the potential site in this phase resource that has potential to become ore is characterized and limited mining methods are considered to aid in a coarse valuation of the prospect at the conclusion of this phase a limited number of feasible alternatives for exploiting the opportunity should be identified and adequately framed for further evaluation in the selection phase selection the reserve determination from the identification phase is the basis for semiquantitative mine plan comparisons competing mine design alternatives are compared in pro forma economic evaluations and investment performance measures such as net present value along with scored risk assessments uncertainty that leads to variability of outcomes risk will be characterized and mitigating

strategies or controls will be developed should the decision to move into the implementation phase be approved the preferred mine design in terms of financial value and technical feasibility results from this stage of planning at the conclusion of this phase a single preferred alternative for the mine plan should be selected for optimization in the definition phase definition in this phase all of the detailed planning and resource estimation of the prior phases are refined and optimized to deliver a final plan prior to implementation the success of this phase will define the success or failure of the venture gaps in information inaccurate planning or even human resource failures can lead to loss of investment environmental damage human injury and negative community impacts the key to success in execution is to invest in front end planning and design prior to implementation which should follow a rigorous plan that includes sufficient contingency and flexibility to manage the variability that is inevitably encountered room and pillar mining method the r p mining method is a popular mining method for underground mining in tabular and lenticular deposits as shown in figure 6 4 1 it is the dominant choice for noncoal underground mining and is frequently applied in coal mines the concept is to sink a shaft or construct a slope or drift depending on depth of ore to the elevation of the mining horizon and begin excavating the ore laterally within the deposit where drilling and blasting are not required the focus of the operation is the continuous miner figure 6 4 2 which utilizes a large rotating drum to break the material in front of it an internal gathering system then loads the broken ore onto an onboard conveyor the onboard conveyor feeds onto a shuttle car or articulated hauler which takes the product to an optional mobile belt feeder if present the feeder meters the product onto a conveyor belt which in turn carries the ore to the surface alternatives to shuttle cars and rubber tired haulers batteryor diesel powered are generally termed continuous haulage systems and include bridge conveyors composed of multiple independent bridge carriers and flexible conveyor trains using a single continuous belt mounted on a mobile base frame with bends to follow a producing machine roof support is an integral part of the mining process and is usually done with roof bolts and their relative the roof truss in place change continuous mining the continuous miner makes a cut and roof bolts are installed with a mobile machine called a roof bolter of course ventilation and face drainage are requirements of any mining method alternatively in very weak roof conditions continuous miners with roof bolting equipment onboard are common these machines called bolter miners cut and load the ore simultaneously with the installation of roof bolts this results in in place mining in contrast to the place changing method anothe

r variation of continuous miner has cutting rotors that rotate parallel to the working face these borer miners are popular in potash and trona mining because of their weight and power the inability to adequately ventilate methane in the face with borer miners has forced their decline in coal mining broken ore haulage can be the same for any of the continuous mining methods although the combination of in place mining with continuous haulage methods has some advantages in the r p method a continuous miner excavates the deposit in a grid like pattern driving entries rooms approximately 15 20 ft wide at the intended mining height these openings run parallel to each other along the long axis of the workings crosscuts driven in the same manner at an acute angle to the entries connect the entries to complete the gridlike pattern pillars are left behind to support the roof hence the term room and pillar alternately know as bord and pillar the optimal or favorable characteristics for r p mining are shown in table 6 4 6 this mining method is optimal for minerals with lower ore strength such as coal potash salt or trona r p mining can be practiced with partial extraction to leave behind larger pillars lower resource recovery and higher cost where concerns exist over ground stability or surface subsidence alternatively high extraction mining can be executed where pillar recovery is done after initial panel development this method is productive and cost effective but has rising concerns associated with ground control during the final phases of pillar recovery mobile roof supports have been introduced to help mitigate roof control concerns and reduce logistics related to roof support although an improvement mobile roof supports do not fully address the concern another variation of the r p method is rib pillar extraction where long narrow pillars are developed and recovered in a progressive process intended to improve safety and productivity this technique is effective where some attribute of the deposit does not lend itself to efficient longwall operation yet another variation of r p mining arguably hybridized with longwall mining is shortwall mining in which a continuous miner works with shuttle cars and specially designed roof supports similar to longwall shields or chocks again this method has only found successful application in a few cases for the most part underground soft rock operators are migrating to longwall mining which has left high extraction r p mining on the decline where alternatives are available the r p mining method has distinct advantages the foremost that with a continuous miner operations are nearly continuous in nature most mining sequences require drilling blasting loading hauling dumping and roof support as well as continuous ventilation and drainage the invention of the continuous miner eliminated the independent steps of drilling blasting and loading which substantially incre

ases the overall efficiency of the method and improves general productivity low operating costs and high production rates are typically associated with modern mechanized r p mining continuous miners can cut through soft rock deposits particularly coal with ease resulting in rapid development rates r p mining is generally more flexible than other methods because continuous miners can move to other working places within a panel or possibly across a mine with limited difficulty also the grid layout of the mine allows for straightforward ventilation with consistent airflow to all working faces the major disadvantage of continuous mining is that it can only be applied to a limited variety of applications a continuous miner cannot operate efficiently if at all in harder rocks like limestone or granite and thus its principal advantages cannot be shared r p has been used in a variety of soft rock applications as well as a few hard rock mining applications but on a small scale when compared to coal to purchase equipment and perform development excavations r p requires a moderate capital investment the method is also limited by depth the pillar size is dictated by the weight of the overburden above the deposit so conceptually the deeper the ore body the larger the pillars must be larger pillars result in lower recoveries and overall mining efficiencies pillars can be recovered after initial development by utilizing retreat mining high extraction the primary advantages and disadvantages of continuous miner based r p mining are summarized in table 6 4 7 the difference between coal and noncoal production methods are four main factors 1 strength higher strength generally correlates with higher specific energy of cutting and lower productivity in cutting applications 2 scale coal mines are usually larger in throughput than other soft rock mines because of the need to economically produce a lower value product 3 methane ch4 where most coal mines are gassy many noncoal mines are free of that hazard in most countries coal mines and their related equipment are governed by strict regulations designed to prevent methane or coal dust explosions large mine explosions are often the product of methane explosions which entrain coal dust in the air leading to subsequent and more energetic coal dust explosions in rapid succession worldwide systems involving water or incombustible dust rock dust are implemented to prevent coal dust explosions spontaneous combustion is also a hazard in many coal mines worldwide and is an attribute of some coal seams but not all 4 coal workers pneumoconiosis more commonly known as black lung this chronic debilitating disease is related to excessive exposure to respirable coal dust usually during employment in coal mining the management of dust in coal mining is subject to strict regulations but continues to be an area of industry and regulatory focus longwall mining method longwal

l mining is combined with r p mining to create some of the most efficient and highest producing underground mines in the world first the main entries are driven with the conventional r p techniques using continuous miners a series of panels branching perpendicular from the mains or submains are outlined by a 2 3 entry r p border leaving a very large solid block panel of ore within its confines typical panel dimensions in contemporary coal mines are 800 to 1 400 ft of face length width with 6 000 to 15 000 ft of panel length in coal panel tonnages of almost 12 1 million metric tons are possible in modern longwall faces a shearer armored face conveyor afc stage loader and line of powered roof supports shields are assembled in a setup room at the beginning of the panel before longwall mining commences figure 6 4 3 utilities used by a longwall include emulsion pumps with capacities of 300 500 gpm at 4 000 to 4 800 psi electrical controls powered by 3 300 to 14 400 v provide power to the longwall face equipment at 1 000 4 160 v the shearer moves back and forth across the coal block excavating 100 of the ore within its height capability causing the material to fall onto the afc and be transported to the main belt conveyor system via the stage loader which normally has an integral crusher to provide suitably sized material for conveyor belts the shields advance sequentially following the shearer to hold up the roof directly above the face equipment and advance the afc to repeat the cutting cycle the excavated area behind the shields is allowed to collapse retreat of the longwall progresses as continuous miners develop additional adjacent longwall panels when the longwall reaches the end of the panel specialized activities are executed by a carefully choreographed plan to withdraw the longwall equipment from the completed panel and reinstall it in the next panel during this process key elements of the equipment are refurbished or exchanged with machine manufacturers for an already refurbished machine or component commonly shearers afc and stage loader components pumps and selected electrical equipment are refurbished as necessary to allow high availability in service during the new panel for world class longwalls production can range from 6 6 to 13 2 million tons per year with unplanned production outages resulting in lost opportunity costs estimated to range from 200 to 1 000 min a sketch of the longwall mining method and the components used are shown in figure 6 4 4 in coal seams lower than 60 in thick plow type longwalls are sometimes applied these systems do not have the productivity of higher height shearer based systems and are more vulnerable to abnormal geologic conditions or roof falls on the face or in the tailgate however they are a viable alternative if mining heights below the limits of shearers are required alternatively interest is emerging in mining very thick seams more than 18 ft

in height by longwall methods this has led to development of some very large single pass longwall systems multilift longwalls with limited success and top coal caving longwalls which seem to offer good potential although both retreating and advancing style longwall systems have been used in the past most installations worldwide are now retreating faces this choice causes higher initial development but minimizes the huge task of maintaining gate roads in the caved area behind the face gob goaf the largest number of longwalls is composed of dated equipment styles including low capacity chock type roof supports or even the earliest style prop and bar or timber roof support these are notable only because many such installations still exist worldwide but are clearly inferior to modern technology from productivity and safety perspectives it is of passing interest that a variation of the longwall method is also applied to hard rock gold and platinum reef deposits in southern africa there drilling and blasting break the rock and low production conveying systems and slushers clear broken ore from the face nonexplosive rock breaking or cutting is being evaluated but is not yet commercialized not all soft rock ores are suited for longwall mining which works best in deposits that are laterally extensive flat lying of fairly uniform thickness and primarily free of discontinuities such as faults coal beds deeper than 1 000 ft usually must be extracted by way of longwall mining because using r p methods would require the use of much larger pillars to support the roof and thus reduce the amount of coal that can essentially be recovered rock bursts mountain bumps and outbursts are all manifestations of stored energy release where r p or longwall mining has been conducted with some combination of the following conditions present depth greater than 1 500 ft strong ore or stiff strong rock members in the nearby underlying or overlying strata unexpectedly high in situ horizontal stresses or stress increases from interaction between workings substantial reservoir or pore pressures of pressurized fluids particularly co2 or ch4 these events can range from mild thumps of little significance to catastrophic events capable of serious equipment damage and fatal injury to personnel expert assistance should be enlisted to assist the mine planning process where such events may occur the optimal characteristics for longwall mining are shown in table 6 4 8 even more so than r p mining the longwall method is exceptionally efficient and has outstanding production rates and low operating costs the operation is almost completely mechanized and recovers an extremely high percentage of the ore body the working face is also safe since the roof is directly supported at all times by heavy duty shields electronic controls and automation allow personnel to position themselves away from most of the recognized hazards if conditions allow long

wall mining is the most effective way to excavate a thin tabular deposit with lower ore strength over the years significant improvements to longwall mining equipment have been made to help yield higher production rates shields with high yield capacities and electro hydraulic controls have replaced manually operated frames and chocks afcs have become more robust and powerful with increased size and speed of the chains allowing higher conveyor capacities the shearers have also become much more powerful and reliable which enables more production less downtime and greater equipment longevity there are however a few disadvantages to longwall mining first it requires a substantial capital investment to purchase the highly specialized equipment to create a longwall section the development time is significant because the continuous miners have to progress the main entries and develop gate roads for the longwall panel before the longwall can be installed finally there is little selectivity or flexibility after mining commences longwall mines should be large longlived operations with high production rates in order to make certain of an adequate return on the mine operator s investment the primary advantages and disadvantages of longwall mining are summarized in table 6 4 9 small scale mining methods typically in small scale mining operations a more traditional mining method is favored where pneumatic drills are used to drill holes to be charged with explosives and the ore is then blasted and hauled away small scale coal mines may use this method because access to capital is difficult and the cost of equipment for a continuous miner section is prohibitive where conditions and capital availability permit some operators employ continuous miners in small r p mines the evolving regulatory and socioeconomic climate is likely to systematically diminish such small operators in preference for larger scale operations conclusions the process of selecting the optimum mining method for a given deposit is complex and requires extensive collection of geological metallurgical and mining related data in addition to the analysis of multiple alternatives a thorough understanding of the sociopolitical setting pertinent environmental concerns and applicable regulations is critically important this chapter has discussed the primary deposit characteristics and the mining method performance variables that are involved when selecting a mining method for soft rock extraction because mineable ore deposits exist in all shapes and sizes and no two are alike the best method selection process is not always evident however several key tasks should always be undertaken during method selection for any ore deposit the first step is to identify the mineral resource available the second step is to match the most suitable mining method to the ore body as part of this step it is important to identify pertinent economic or environmental factors that

may constrain methods selection above all else it cannot be overemphasized that mine planners must value the principle of zero harm which encompasses health safety environment and community impacts failure in any of these areas can affect the sustainability of a mining operation just as seriously as a planning or execution failure a review of numerous case histories of success and failure highlights the fact that failed projects are usually due to inadequate deposit characterization inadequate risk assessment and consequent acceptance of elevated risk inadequate planning or overestimation of operating performance or inadequate capital to correctly implement plans there are many paths to failure and the paths to success are few and normally difficult an overview of the planning required for an underground mine is necessarily complicated by the availability of many different types of underground mining methods for a detailed description of these mining and development methods and associated case studies see hustrulid and bullock 2001 the major objective when deciding upon and planning a mining method should be to maximize value which is achieved after consideration of the following r l grayson personal communication safety of all personnel lowest production cost per metric ton maximum productivity required quality and quantity of the final product maximum recovery of reserves optimal environmental considerations there are obviously trade offs in trying to reach these nonaligned objectives room and pillar mining method room and pillar r p mining is a system where a series of rooms horizontal openings are extracted leaving ore rock or coal called pillars in place between the rooms these rock ore coal pillars can be smaller horizontally than the rooms typical for hard rock mining see figure 13 1 5 in chapter 13 1 or larger than the rooms typical for soft rock or coal mining normally rooms and pillars are of consistent size and shape and are laid out in a uniform pattern however for some metal mining because the ore is not uniform pillars can be varied in size and location to enable placement in lowgrade areas of the stope this is sometimes called stope andpillar mining the term room and pillar mining has been applied to metal mining in the united states for more than 150 years in the vast mining districts of the missouri lead belts the tennessee zinc district and the tri state zinc district of southwestern missouri southeastern kansas and northeastern oklahoma considering the number of underground mines of coal dolomite gypsum limestone potash salt and trona as well as all of the mississippi valley type lead and zinc mines it should not be surprising to realize that approximately 60 to 70 of all underground mining in the united states is some form of r p mining this amounts to nearly 340 mt a 370 million tons yr zipf 2001 for the aggregate industry alone according to niosh the

re are 90 to 100 underground mines all of them r p iannacchione 1999 in today s permitting environment at any given time there are probably between 20 and 40 r p underground aggregate mines being planned access to the r p mine hard rock and coal mining methods although the access to a mine is not always influenced by the mining method some discussion is warranted on the various approaches to the initial mine and production opening where r p mining is to be applied if it is possible to develop the resource from a hillside adit doing so obviously provides the least expensive and most complete method of entry in coal mining this is called a drift entry if a shaft is sunk bullock 1973 it should be sunk somewhere close to the center of gravity of the ore body unless the ground is going to be allowed to cave in which case it should be placed well outside the cone of subsidence sunk to a depth that allows most of the ore that is be hauled downgrade to reach the shaft storage pockets sunk deep enough to accommodate adequate storage pockets skip loading and a crusher station if needed and located for aesthetic reasons such that the headframe is out of sight of the public if a decline is to be driven the maximum grade of the decline depends on the equipment that will be driven on the decline for truck haulage the decline gradient should be matched to the gearing and optional power train of the trucks here are a few planning guidelines depending on how a decline will be used for truck haulage for trackless haulage 12 5 is the maximum grade recommended for normal mine trucks if super powered trucks are planned then grades up to 17 may be acceptable for conveyor belt haulage where rubber tired trackless equipment must negotiate on a regular basis 15 is the maximum grade recommended for conveyorbelt haulage only the theoretical maximum grade is approximately 0 17 to 0 31 radians 10 to 18 depending on the type of material cema 1994 beyond these recommended angles of incline material will slide down the belt en masse and internally on top of itself and lumps will roll down the belt and over the top of the fines however equipment must be able to access the belt to occasionally clean up spill rock unless hand shoveling for cleanup is planned room and pillar stoping for hard rock the differences in strength hardness and abrasiveness of rocks such as limestone dolomite or sandstone as compared to those of the soft materials coal potash salt or trona necessitate different extraction methods primary extraction methods one advantage of modern r p mining systems is that every task can be mechanized to some degree provided that it is economically sound to do so mechanization minimizes the operating labor force and simplifies staffing the high capacity equipment for modern r p operations is relatively simple to operate although most r p mining is done by drilling and blasting particul

arly for aggregates and metals some mining is done by mechanical excavation usually with roadheaders bullock 1994 with the power of today s mechanical excavating machines and with improvements that are being made in tools such as disk and pick cutters the possibility of mechanical excavation should be at least considered during the feasibility study for any rock under 100 mpa 15 000 psi or even up to 136 mpa 20 000 psi if it contains fractures and is low in silica content where mechanical excavation is truly viable its use equates to higher production rates and reduced operating cost the following are advantages of mechanical excavation where it is viable ozdemir 1990 improved personal safety minimal ground disturbance reduced ground support needed continuous noncyclic operations low ground vibrations and no air blast uniform muck size less crushing and grinding in the mill reduced ventilation requirements conducive to automation room width for productivity reasons rooms should be as wide as is practical and safe the wider the rooms the more efficient the drilling and blasting and the larger and more efficient the loading and hauling equipment however room width is always limited by the rock mass strength of the ore body back and floor compared to the stress levels induced into the rock it is inappropriate to design room widths simply from elastic theory without taking into account rock mass strength however since rock and pillars can be reinforced to increase the effective rockmass strength final room width may be a matter of economics many discussions have been written on how to design a roof span for further insight into the rock mechanics of r p roof spans see hustrulid and bullock 2001 what is important at this point is to determine what information is needed for the design and how much of the needed information is already at hand chapter 12 1 contains a general summary of geological and structural information that should have been determined during exploration of an ore body unfortunately most exploration groups spend little time or money in determining the information that is needed to construct a rock mass classification of the mineralized areas and rock surrounding the mineralization a best guess rock mass analysis may have to be done with nothing but the exploration information in any case it is hoped that mapping of underground structures from core logs surface mapping possibly mapping of surface outcrops of the same underground structures and geophysical information along with the rock quality designation rqd of the rock core are sufficient for a crude rock mass classification to be constructed pillar width the overall strength of a pillar is related to its height that is the ratio of pillar width w to pillar height h is important the amount of load that a pillar can safely carry is proportional to the ratio w h thus a pillar of ratio 4 1 has a much larger safety fac

tor than does a pillar of ratio 1 1 or 1 5 ratios of 1 3 to 1 4 are not uncommon in some competent hard rock metal mines the theoretical load as calculated by the overburden load distributed to the pillars may or may not be the load that is actually carried there is a good chance that the load may arch over some of the interior pillars of the stope and transfer load to barrier pillars or waste areas in some such cases interior pillars can be made smaller as yielding pillars if a stopes mine is very wide a row of large rectangular barrier pillars should be retained at regular intervals in areas of very large lateral extent this prevents cascading pillar failure of the entire area in a domino affect zipf and mark 1997 for more information the reader is referred to case studies on r p stoping in hustrulid and bullock 2001 these studies describe how different mines approach these design problems as well as catastrophic failures that have occurred when proper precautions were not taken secondary extraction methods pillar removal should be planned as part of the overall mining of areas where the economic value of what remains warrants the extraction of some or all pillars for example it is not uncommon for some very high grade pillars in the lead zinc copper mines of the viburnum trend in missouri to have a value of more than 1 million per pillar for optimum recovery the initial pillar design must be correct and include barrier pillars to prevent catastrophic failures there are five basic approaches to removing pillar ore after completion of primary mining 1 partially extract by slabbing the highest grade part of each pillar 2 remove a certain number of the high grade pillars completely but leave enough to support the back 3 encapsulate pillars with fenced cemented rock fill or paste backfill then come in underneath the pillar and drop it into the sublevel below 4 encapsulate low grade pillars with fenced cemented rock fill or paste backfill to form a barrier pillar then remove the surrounding pillars lane et al 1999 2001 5 for narrow areas and strong reinforced back remove all pillars room and pillar mining for coal and soft rock for the typical r p mine layout for mining of bituminous coal shown in figure 6 5 1 five main entries allow access to the production panel through panel entries coal mine dimensions are generally as follows entry widths in the united states can be up to 6 m 20 ft and are generally driven 18 3 to 30 5 m 60 to 100 ft apart center to center panel widths can be 122 to 183 m 400 to 600 ft limited primarily by the cable reach of the electric shuttle cars that are usually used to move the coal panel lengths generally vary from 610 to 1 220 m 2 000 to 4 000 ft but can be longer an overview of the planning required for an underground mine is necessarily complicated by the availability of many different types of underground mining methods for a detailed description of t

hese mining and development methods and associated case studies see hustrulid and bullock 2001 the major objective when deciding upon and planning a mining method should be to maximize value which is achieved after consideration of the following r l grayson personal communication safety of all personnel lowest production cost per metric ton maximum productivity required quality and quantity of the final product maximum recovery of reserves optimal environmental considerations there are obviously trade offs in trying to reach these nonaligned objectives room and pillar mining method room and pillar r p mining is a system where a series of rooms horizontal openings are extracted leaving ore rock or coal called pillars in place between the rooms these rock ore coal pillars can be smaller horizontally than the rooms typical for hard rock mining see figure 13 1 5 in chapter 13 1 or larger than the rooms typical for soft rock or coal mining normally rooms and pillars are of consistent size and shape and are laid out in a uniform pattern however for some metal mining because the ore is not uniform pillars can be varied in size and location to enable placement in lowgrade areas of the stope this is sometimes called stope andpillar mining the term room and pillar mining has been applied to metal mining in the united states for more than 150 years in the vast mining districts of the missouri lead belts the tennessee zinc district and the tri state zinc district of southwestern missouri southeastern kansas and northeastern oklahoma considering the number of underground mines of coal dolomite gypsum limestone potash salt and trona as well as all of the mississippi valley type lead and zinc mines it should not be surprising to realize that approximately 60 to 70 of all underground mining in the united states is some form of r p mining this amounts to nearly 340 mt a 370 million tons yr zipf 2001 for the aggregate industry alone according to niosh there are 90 to 100 underground mines all of them r p iannacchione 1999 in today s permitting environment at any given time there are probably between 20 and 40 r p underground aggregate mines being planned access to the r p mine hard rock and coal mining methods although the access to a mine is not always influenced by the mining method some discussion is warranted on the various approaches to the initial mine and production opening where r p mining is to be applied if it is possible to develop the resource from a hillside adit doing so obviously provides the least expensive and most complete method of entry in coal mining this is called a drift entry if a shaft is sunk bullock 1973 it should be sunk somewhere close to the center of gravity of the ore body unless the ground is going to be allowed to cave in which case it should be placed well outside the cone of subsidence sunk to a depth that allows most of the ore that is be hau

led downgrade to reach the shaft storage pockets sunk deep enough to accommodate adequate storage pockets skip loading and a crusher station if needed and located for aesthetic reasons such that the headframe is out of sight of the public if a decline is to be driven the maximum grade of the decline depends on the equipment that will be driven on the decline for truck haulage the decline gradient should be matched to the gearing and optional power train of the trucks here are a few planning guidelines depending on how a decline will be used for truck haulage for trackless haulage 12 5 is the maximum grade recommended for normal mine trucks if super powered trucks are planned then grades up to 17 may be acceptable for conveyor belt haulage where rubber tired trackless equipment must negotiate on a regular basis 15 is the maximum grade recommended for conveyorbelt haulage only the theoretical maximum grade is approximately 0 17 to 0 31 radians 10 to 18 depending on the type of material cema 1994 beyond these recommended angles of incline material will slide down the belt en masse and internally on top of itself and lumps will roll down the belt and over the top of the fines however equipment must be able to access the belt to occasionally clean up spill rock unless hand shoveling for cleanup is planned room and pillar stoping for hard rock the differences in strength hardness and abrasiveness of rocks such as limestone dolomite or sandstone as compared to those of the soft materials coal potash salt or trona necessitate different extraction methods primary extraction methods one advantage of modern r p mining systems is that every task can be mechanized to some degree provided that it is economically sound to do so mechanization minimizes the operating labor force and simplifies staffing the high capacity equipment for modern r p operations is relatively simple to operate although most r p mining is done by drilling and blasting particularly for aggregates and metals some mining is done by mechanical excavation usually with roadheaders bullock 1994 with the power of today s mechanical excavating machines and with improvements that are being made in tools such as disk and pick cutters the possibility of mechanical excavation should be at least considered during the feasibility study for any rock under 100 mpa 15 000 psi or even up to 136 mpa 20 000 psi if it contains fractures and is low in silica content where mechanical excavation is truly viable its use equates to higher production rates and reduced operating cost the following are advantages of mechanical excavation where it is viable ozdemir 1990 improved personal safety minimal ground disturbance reduced ground support needed continuous noncyclic operations low ground vibrations and no air blast uniform muck size less crushing and grinding in the mill reduced ventilation requirements conducive to au

tomation room width for productivity reasons rooms should be as wide as is practical and safe the wider the rooms the more efficient the drilling and blasting and the larger and more efficient the loading and hauling equipment however room width is always limited by the rock mass strength of the ore body back and floor compared to the stress levels induced into the rock it is inappropriate to design room widths simply from elastic theory without taking into account rock mass strength however since rock and pillars can be reinforced to increase the effective rockmass strength final room width may be a matter of economics many discussions have been written on how to design a roof span for further insight into the rock mechanics of r p roof spans see hustrulid and bullock 2001 what is important at this point is to determine what information is needed for the design and how much of the needed information is already at hand chapter 12 1 contains a general summary of geological and structural information that should have been determined during exploration of an ore body unfortunately most exploration groups spend little time or money in determining the information that is needed to construct a rock mass classification of the mineralized areas and rock surrounding the mineralization a best guess rock mass analysis may have to be done with nothing but the exploration information in any case it is hoped that mapping of underground structures from core logs surface mapping possibly mapping of surface outcrops of the same underground structures and geophysical information along with the rock quality designation rqd of the rock core are sufficient for a crude rock mass classification to be constructed pillar width the overall strength of a pillar is related to its height that is the ratio of pillar width w to pillar height h is important the amount of load that a pillar can safely carry is proportional to the ratio w h thus a pillar of ratio 4 1 has a much larger safety factor than does a pillar of ratio 1 1 or 1 5 ratios of 1 3 to 1 4 are not uncommon in some competent hard rock metal mines the theoretical load as calculated by the overburden load distributed to the pillars may or may not be the load that is actually carried there is a good chance that the load may arch over some of the interior pillars of the stope and transfer load to barrier pillars or waste areas in some such cases interior pillars can be made smaller as yielding pillars if a stopes mine is very wide a row of large rectangular barrier pillars should be retained at regular intervals in areas of very large lateral extent this prevents cascading pillar failure of the entire area in a domino affect zipf and mark 1997 for more information the reader is referred to case studies on r p stoping in hustrulid and bullock 2001 these studies describe how different mines approach these design problems as well as catastrophic failures that have occurre

d when proper precautions were not taken secondary extraction methods pillar removal should be planned as part of the overall mining of areas where the economic value of what remains warrants the extraction of some or all pillars for example it is not uncommon for some very high grade pillars in the lead zinc copper mines of the viburnum trend in missouri to have a value of more than 1 million per pillar for optimum recovery the initial pillar design must be correct and include barrier pillars to prevent catastrophic failures there are five basic approaches to removing pillar ore after completion of primary mining 1 partially extract by slabbing the highest grade part of each pillar 2 remove a certain number of the high grade pillars completely but leave enough to support the back 3 encapsulate pillars with fenced cemented rock fill or paste backfill then come in underneath the pillar and drop it into the sublevel below 4 encapsulate low grade pillars with fenced cemented rock fill or paste backfill to form a barrier pillar then remove the surrounding pillars lane et al 1999 2001 5 for narrow areas and strong reinforced back remove all pillars room and pillar mining for coal and soft rock for the typical r p mine layout for mining of bituminous coal shown in figure 6 5 1 five main entries allow access to the production panel through panel entries coal mine dimensions are generally as follows entry widths in the united states can be up to 6 m 20 ft and are generally driven 18 3 to 30 5 m 60 to 100 ft apart center to center panel widths can be 122 to 183 m 400 to 600 ft limited primarily by the cable reach of the electric shuttle cars that are usually used to move the coal panel lengths generally vary from 610 to 1 220 m 2 000 to 4 000 ft but can be longer in figures 6 5 1 and 6 5 2 note that the panel pillars are being mined in the united states this is called pillaring or caving elsewhere it is called stoping normal practice in pillaring is to drive rooms and crosscuts upon advance mining into the virgin coal seam in a panel and to pillar upon retreat mining while moving back out of the panel in the united states the caved area is called the gob elsewhere it is called the goaf pillars are not removed if the surface must be supported pillaring is often omitted for other reasons as well hartman and mutmansky 2002 mining methods two methods of r p coal mining exist conventional and continuous conventional operation involving drilling and blasting of the undercut coal seam is an antiquated method practiced in 5 of coal mines in the united states conventional operation using conventional equipment is cyclic with the following sequence 1 undercut a kerf of 127 to 178 mm 5 to 7 in to improve coal breakage during light blasting 2 drill blastholes with an electric rotary drag bit auger drill 3 charge the holes lightly with permissible explosive and then blast 4 load the coal

with an electric gathering arm loader into a shuttle car for transport away from the face auxiliary operations of roof control ventilation installation and cleanup must also be performed a typical mining section that uses this system is shown in figure 6 5 3 which shows a seven entry development the numbers represent the sequence of faces where each crew member performs their function of the cycle continuous operation is so named because a mechanical excavating drum miner continuously extracts the coal coal is mined continuously and then hauled from the face usually by electric shuttle car but sometimes by conveyor belt or diesel scoop machine there is usually some waiting between the loading of each shuttle car or scoop as with conventional operation auxiliary operations of roof control ventilation installation and cleanup must also be performed the plan view in figure 6 5 2 shows section entries room entries and rooms associated with a production panel within a mine room entries rooms and associated crosscuts are mined upon advance and pillars are mined upon retreat in the figure note the numbering on the entries in the section entry set retreat methods retreat mining methods vary greatly in the coal industry the method used can depend on seam height as well as local miningconditions availability of mining equipment pillar size success of similar methods in adjacent mines opinions of mine engineers and state and federal roof control specialists pillar extraction methods widely practiced in the industry include christmas tree outside lift split and fender and combinations thereof as shown in figure 6 5 4 in a study by the u s national institute for occupational safety and health niosh mark et al 2003 it was estimated that nationally about 60 of retreat mining used the christmas tree method 35 used the outside lift method and only 5 used the split and fender method the christmas tree method is usually favored because it does not require placing changes and bolting using 150 case histories niosh has developed a methodology that has proved successful in generating and validating the analysis of retreat mining pillar stability armps method and program developed by mark and chase 1997 the method aids in the correct design of retreat mining and helps show potential problems such as bump seismic prone areas chase et al 2002 an excellent review of all current practices in the united states has been written by mark 2009 the christmas tree method figure 6 5 4a is also called left right or twinning cuts are taken both left and right on both sides of the entry a continuous miner removes most of the coal on each side until a chevron type pillar typically a corner wedge shaped remnant pillar is left the figure shows a common sequence in which lifts are extracted during barrier and production pillar extraction using mobile roof support mrs units the outside lift method figure 6 5 4b

is the original method developed for use with mrs units this method is suitable for narrow pillars when combined with extended cut mining there are many variations depending on conditions pillar dimensions and coal haulage equipment generally the pillar is sized so that lifts taken from one side of the pillar are sufficient to extract the pillar without going beyond the supported roof cuts are taken starting from the pillar near the gob goaf and moving toward solid coal the sequence of cuts shown in the figure is typical and mrs units are moved in a retreat mining process or adjacent entry with supplemental support generally provided by posts mechanical chocks or mrs units multiple pillars are usually extracted simultaneously to provide an adequate number of working places and so avoid production delays the split and fender method figure 6 5 4c involves a specific cut sequence for continuous mining equipment as shown by the numbers in the pillars again supplemental support in the form of mrs units is used the disadvantages of mrs units are twofold costs for their initial purchase and for recovery if they are trapped by a rock fall are high and their operating range is usually limited to seams thicker than approximately 1 1 m 42 in typical retreat mining methods using breaker posts or mechanized chocks the vast majority of thin seam mines seams 1 3 m or 52 in in the united states use wooden breaker posts mark et al 2003 in a 2002 survey of mines in southern west virginia united states by niosh mark 2002 the vast majority of thicker seam mines were already using mrs units but of the 54 thin seam mines only 7 13 were using them in thinseam mines a timber plan that requires an adequate number of posts installed at the proper times and locations is essential more than 100 roadway turn and breaker posts can be required to extract a single pillar where hardwood timber is not abundant mechanical chocks can be used as is often the case in europe australia and south africa typical examples of this application are sited in the following paragraphs in very low i e thin seams where continuous haulage is used the christmas tree and outside lift methods have been used in combination for pillar systems developed with continuous haulage figure 6 5 4d when using mobile bridge conveyors the most common type of continuous haulage crosscuts are driven at 60 angles to facilitate movement of bridges and carriers the parallelogram shaped pillars create a panel configuration that is usually referred to as a herringbone or turkey foot design each mining cycle starts with recovery of the two central pillars blocks 1 and 2 left standing out in the gob by the previous cycle each pillar is extracted by the outside lift method after cutting lifts 1 and 2 in block 1 the continuous miner is maneuvered to cut lifts 3 and 4 in block 2 a variation is to cut the two central pillars using the christmastree

method the extraction sequence removes the left barrier and block 3 followed by the right barrier and block 4 and then the sequence repeats feddock and ma 2006 in the low seam coal mines of appalachia united states many mines still use the four basic pillar extraction cut sequences described by kauffman et al 1981 and shown in figure 6 5 5 split and fender pocket and wing open ending and outside lift the numbers in the figure represent the sequence of cuts made by a continuous miner temporary prop supports are shown as rows of black dots there are several disadvantages to using posts rather than mrs units mark et al 2003 setting posts exposes miners to ground falls posts have a limited load bearing capacity a typical hardwood post 152 mm 6 in in diameter can carry about 45 t 50 st posts have limited roof convergence range from 25 to 51 mm 1 to 2 in before they break because of their weight and bulk posts can cause material handling injuries particularly in high coal sublevel open stoping mining method r p methods are often used for subhorizontal ore bodies of relatively uniform thickness as strata dip and or ore body thickness increases other extraction methods are usually used for example a moderately thick flat dipping deposit would normally be mined by r p methods but for a dip of 90 loading on the pillars is from the horizontal direction and blasted ore falls down for collection at the bottom of the stope although the general geometry is the same as that for r p methods the generic name given to this system is sublevel open stoping blasthole stoping vertical crater retreat mining and vein mining fall under this general heading a special form of sublevel open stoping is called shrinkage stoping in general the method is applied to ore bodies having dips greater than the angle of repose of broken material 50 so that material is transported to the collection points by gravity for massive deposits stopes with vertical walls are created and the overall dip of the deposit is immaterial the criterion for applying the method is that during extraction the openings created must remain open after extraction the openings may be filled or left open and the pillars left between stopes may be extracted or left in place this section presents some typical layouts used for extracting ore it is assumed that mobile equipment is used with ramp access throughout planning it is important to be aware of the potential for ore dilution or ore loss when drilling long holes and when mining undulating veins in feasibility studies the value of dilution for sublevel long hole stoping should be 15 and ore loss 5 however for short blasthole methods such as shrinkage methods dilution and ore loss might be 5 extraction principles one can consider an ore block of width w 1 to 10 m 3 3 to 33 ft length l 10 to 40 m 33 to 131 ft and height h 20 to 30 m 66 to 98 ft for simplicity

it can be assumed that the block is vertical although for this method the dip of the block is immaterial since if the block is thick enough a stope can be developed that will flow by gravity and can be mined by a number of sublevel stoping techniques blasted ore falls to the bottom of the block and is removed with load haul dump lhd equipment there are various designs for the extraction level it can also be assumed that a trough is created using fans blasted toward an opening slot the lhd units travel in a footwall haulage drift running parallel to the trough access to the trough is from the side the location and number of access drawpoints are such as to provide full extraction coverage with the availability of remote or teleremote operated lhd units it is now common to design stopes with flat bottoms rather than trough undercuts final cleanup of up to 20 of the stope ore is done by remote control which simplifies stope layout and makes drilling and blasting more efficient with remote lhd units there may be no requirement for a footwall haulage and crosscuts because the stope is accessed along the strike within the ore reducing waste development costs blasthole stoping end slicing because it is the simplest and usually lowest in cost blasthole stoping also called end slicing is the first method considered for mining a block figure 6 5 6 from the drilling level at the top of the block rows of parallel blastholes are drilled down to the top of the extraction trough a raise is driven at one end of the block and is slashed to full stoping width to form a slot holes are blasted one or several rows at a time toward the open slot the blasting design and layout is similar to that used in bench blasting hole diameters vary widely but are typically in the range 76 to 165 mm 3 to 6 5 in for wide blocks diameters of 165 mm 6 5 in are often used hole straightness is an important design consideration that affects fragmentation ore loss and dilution in general the largest hole diameter possible for the stope geometry is selected because straight hole length is strongly dependent on hole diameter the amount of development required to exploit a certain volume of ore is inversely proportional to block height and because the cost of development is significantly higher than that of stoping the tallest extraction blocks possible should be planned for the stope sublevel stoping if geomechanic studies indicate that very tall blocks those whose height exceeds the straight drilling length from one drill location can be extracted using the same extraction level then several drilling levels at various heights within the block can be created by a method called sublevel stoping figure 6 5 7 the layout is similar to that for blasthole stoping with an extraction level and an opening slot but now there are multiple drilling levels mining is done either overhand lower drilling blocks are extracted first or underha

nd upper drilling blocks are extracted first overhand stoping is usually assumed the simplest approach is to repeat the drilling layout for one level blasthole stoping ore body thickness is assumed to be such that the full width is undercut and becomes available for drilling access in which case parallel holes can be drilled an alternative approach is to drill fans of holes rather than parallel holes from the sublevels figure 6 5 8 there can be one or multiple drill drifts on each sublevel and the rings can be drilled downward upward or in full rings there are many variations of drill pattern based on a number of factors for example the distance between sublevels needs to be maximized so as to minimize the amount of development per ton of ore broken but the accuracy of the drilling equipment in that particular rock determines how long holes can be without excessive drill hole deviation which causes boulders ore loss ore dilution or so called misbreaks rock that breaks incorrectly or not at all in blasting likewise the width of the ore block to be taken can determine whether sublevel drifts are placed on the hanging wall and the footwall or if the ore is narrow whether a single drift in the ore might approach the limits of the ore another variable is hanging wall strength if a hanging wall needs support then some of the sublevels need to be on the hanging wall side so that sublevel drifts can also be used in place of cable bolts for support as well as for drilling long blastholes yet another variable is whether a primary stope or a pillar between two previously mined filled stopes is being mined in the latter case the blastholes break more easily to the free face of the rock fill interface for a guide to these planning decisions see chapter 13 4 use of sublevel stoping assumes good stability of the openings created stability surprises can mean partial or even full collapse of partially extracted stopes production can be stopped fully because of the presence of large blocks in the drawpoints even in the best case scenario there is likely to be some ore loss and dilution the footwall hanging wall and roof can be reinforced before or during mining extraction blocks stopes can be oriented parallel or perpendicular to the ore body depending primarily on the width of the ore body and how the mine planners want to drive the access drifts to support extraction of the stope block vertical crater retreat stoping in the cases discussed previously rings of holes are blasted toward a vertical slot in vertical crater retreat vcr or vertical retreat mining vrm systems the need for a slot connecting the drilling and extraction level has been eliminated simplifying development figure 6 5 9 rather the slot is replaced by a horizontal slot undercut created at the bottom of the block on the extraction level a real trough can be created but is not necessary from the drilling level large diameter 165 m

m or 6 5 in parallel holes are drilled downward to the undercut level short explosive charges length 6 the hole diameter are lowered to positions slightly above the top of the undercut these spherical charges are detonated dislodging a crater or cone shaped volume of rock into the underlying void as each layer of charges is placed and detonated mining of the stope retreats vertically upward in a vcr the design of the blasting pattern is based on full coverage of the block cross section by the adjacent craters normally the blasting pattern is tighter holes are spaced closer than for large hole blasthole stoping and hence the powder factor is larger during blasting under these confined conditions fragmentation is generally finer than it is for blasthole stoping prior to charge placement care must be taken in determining the location of the free surface in some cases ore blocks can overbreak vertically along fractures or faults and leave a vertical opening for the next round to break to in which case the next round in a nearby hole will break sideways to any nearby free face thus observations and measurements are essential prior to planning every blast in this system the level of broken rock remaining in the stope can be controlled to provide varying levels of support to the stope walls if the stope is kept full except for a small slot to provide a free surface and swell volume for the blasted rock in the slice the method used is classified as a shrinkage method and the remaining ore is drawn out at the completion of mining vein mining another approach to extracting the ore block is called vein mining figure 6 5 10 at the highest level of the block to be extracted a connection is made to the ore body access is assumed to be via the footwall side and connection is made in the middle of the extraction block on the extraction level an undercut or extraction trough is prepared a raise is driven between the extraction level and the upper access point from which long blastholes can be drilled using a boliden type cage or the alimak technique ovanic 2001 a raise is driven upward in the footwall a small distance from the ore footwall contact the next step in the process is the drilling of subhorizontal fans of blastholes using the alimak platform or bolidentype cage in such a way that the plan area of the extraction block is fully covered the hole diameter is determined by the capacity of the drilling machine but should be as large as possible since the toe spacing and the burden distance between fans is determined by the hole diameter and the explosive used when drilling of the entire extraction block is complete the fans are charged and blasted one or more at a time working off the raise platform access to the block is now only from the upper level as the stope is retreated upward the ore in the stope can be removed after each blast or can be left in place and only enough removed to provide swell volum

e for the next one or more slices rock reinforcement can be installed in the hanging wall if required from the raise platform during the drilling of the production holes an advantage of this method is that it enables extraction of very high ore blocks with a minimum of development upper access point extraction level and connecting raise the overall length of the extraction block is determined by the straight hole drilling length of the available drilling equipment however if larger blocks are more economically mined then multiple raises can be used and this possibility should be considered in the feasibility study a disadvantage is that drilling and charging must be done from a raise environment which can be hazardous unless well managed however major advances have been made in the mechanization and automation of rigs used for drilling shrinkage stoping although normally considered as a separate method it is logical to also discuss shrinkage stoping at this point since it is an open stope method figure 6 5 11 the method is generally applied to very narrow extraction blocks that have traditionally not lent themselves to a high degree of mechanization it has been applied successfully in high grade precious metal mining because of its low dilution and low ore loss the extraction block is laid out longitudinally due to the very narrow nature of the ore body recovered an extraction drift is located in the footwall with loading crosscuts positioned at regular intervals raises are driven at each end of the extraction block connecting to the above lying level an initial horizontal extraction slice is driven across the block from raise to raise extraction troughs are created by drilling and blasting the rock between this level and the underlying extraction points when the extraction system has been created short vertical holes are drilled into the roof of the first extraction slice using the raise access the miners stand on the broken ore that forms the working floor jackleg or stoper drills are used to drill small diameter holes the holes are charged and then ore is extracted from the stope to provide room for the blasted material the blast is initiated and the miners reenter the newly created void to bar down loose material before drilling out for the next slice the process continues working upward one slice at a time when the upper end of the extraction block is reached the ore is drawn out until that time the stope is filled with broken ore a severe disadvantage of this method is that it is among the most hazardous of mining methods the miner is exposed to freshly blasted unsupported ore while working on rough broken ore in addition voids can be created in the ore being drawn which can suddenly collapse in areas where miners are working for these reasons as well as the relative high operating cost its use is limited in the united states however it is still used fairly extensively in developing countries a

nd in canada summary of open stoping sublevel mining depending on the geometry of the ore body several varieties of sublevel stoping can be used however the ore bodies must all have strong wall rocks and competent ore either naturally or by means of reinforcement since in the process of ore removal large openings are created the extraction block used to illustrate the layouts for the different mining systems can now be duplicated and translated laterally and vertically in the ore body leaving pillars to separate adjacent blocks the size and shape of the extraction block can be adjusted to fit the ore body geometry and the mine infrastructure openings created during this primary mining can be filled with various materials or left unfilled the filling materials can be cemented or left uncemented depending on the next stage of recovery envisioned various methods can be used to recover remaining reserves tied up in the pillars during the feasibility studies these secondary recovery methods should be examined at the same time as the primary system is designed although for simplicity the basic extraction block was considered to be vertical the process can obviously be repeated for ore bodies having various dip conditions cut and fill mining method where ore and or wall rocks are weak and hence both opening size and allowable time between ore removal and filling of the excavation is strictly limited a number of extraction designs can be applied all of which fall under the general category of cut and fill mining this versatile method can be adapted to the extraction of any ore body shape with some exceptions all of the ore is removed via drifts that are then filled as a result mining costs are higher than for other methods but when the method is applied correctly recovery is high and dilution is generally low thus it is appropriate for the extraction of high grade ore bodies extraction principles for simplicity an extraction block of the same type used in the previous section is assumed access is via a ramp driven in the footwall and mobile equipment is used typically drifts used in mechanized cut and fill mining are about 5 m 16 4 ft high the ore block to be extracted is assumed to be vertical and of a width that can be removed by means of normal drifting when ore body strength is fairly good overhand cut and fill mining is normally applied figure 6 5 12 ideally access to each level is via crosscuts originating at the midlength position of the block such that two headings can be operated at one time typical drift rounds consisting of drilling blasting loading scaling and installation of rock reinforcement are used this progression of operations can lead to delays unless carefully planned drilling of heading 2 is carried out while other operations are being done at heading 1 when the extraction slice is complete backfilling takes place fill is placed so as to leave a small gap to the overlying ore fi

gure 6 5 13a such that a heading 5 m 16 ft high recovers about 4 m 13 ft of ore on the next slice this gap forms the free surface for blasting the process continues upward slice by slice to the top of the block several extraction blocks can be operated simultaneously to meet production requirements the horizontal pillar created between two such stacked extraction blocks is called the crown pillar for the underlying stope and the sill pillar for the stope above normally the first cut of the extraction block above the sill is filled with cemented fill to facilitate later extraction of the pillar in some cases the wall rock is strong enough to allow a double slice to be open at any given time the first slice is mined by drifting and then rather than filling directly upwardoriented drill holes called uppers are drilled the length of the slice figure 6 5 13b when drilling is complete several rows of holes are charged and blasted beginning at the ends of the extraction block and retreating toward the access ore is extracted by an lhd unit after each blast and transported to the orepass efficiency can thus be improved by changing the typical cycle to one in which all drilling is done first followed by charging and loading then either one lift can be backfilled followed by the drilling of uppers or both lifts can be backfilled followed by drifting and then drilling of uppers access to this one drift width cut and fill stoping is via an access ramp in the footwall often where the wall rock is fairly competent four slices are accessed from a given point on the ramp figure 6 5 12 in overhand cut and fill stoping crosscut 1 is made first when the slice is complete the roof of the crosscut is slashed down to form crosscut 2 this continues for the four slices at which time a higher point on the ramp is selected as the origin of the crosscuts generally the maximum crosscut inclination is limited to about 20 although 12 to 15 is more normal this sets the position of the ramp with respect to the ore body there are several variations of overhand cut and fill stoping using so called ramp in stope development this method has the advantage that development is in ore although some development must still be done in the footwall it can be applied to lower grade deposits for which the development cost of conventional mechanized cut and fill mining cannot be justified although overall operations are usually less efficient one such mine development is the bruce copper mine in bagdad arizona united states described by johnson et al 1998 another general description of ramp in stope cut andfill mining is in pugh and rasmussen 1998 if the strength of the wall rock and ore is quite good spans of more then two lifts can be created by means of the avoca rill mining or benching mining method figure 6 5 14 in the figure slices 1 and 4 are extracted by drifting stage 1 rows of vertical blastholes are then dr

illed from the floor of slice 4 to the roof of slice 1 a vertical slot is created and rows of holes are blasted one or more at a time toward the slot ore is extracted by an lhd unit operating in slice 1 at the same time that retreat extraction is under way filling is being conducted from the opposite end of the stope stage 2 a gap is maintained between the extraction and filling fronts to minimize dilution when completed slice 7 is removed by drifting stage 3 slices 5 and 6 are now removed using slice 7 as the drilling level and slice 4 as the extraction level if the extraction block is quite wide the cut and fill method can still be used but now several drifts are driven side by side figure 6 5 15 this is similar to r p rib pillar mining with the rooms being filled and the pillars then being extracted various techniques can be used to shape the drifts the most common is shown in figure 6 5 15c here straight walls are used and every other drift is removed in a primary mining phase cemented fill is used to avoid dilution during removal of the interlying drifts another variation is to make the primary drifts narrow and the secondary drifts wide to minimize the use of cemented fill if the strength of the ore or hanging wall is very poor e g in the range of rock mass rating rmr 20 to 40 which might be the equivalent of a standup time depending on the roof span of 1 to 100 hours then the underhand cut and fill mining method often shortened to undercut and fill can be used figure 6 5 16 the first slice is taken and then various techniques are used to prepare a layer that becomes the roof when the slice below is extracted in the past the most common mining method involved a timber floor pinned into the walls it is now more common to pour a layer of cemented fill with or without reinforcement the remainder upper portion of the drift can be left open or filled with uncemented fill the next slice is then extracted under the constructed roof a more common practice in north america is to use engineered cement or paste fills when they are used for mining wide ore bodies the material must be jammed tight to the back or previous floor from the same development access level some mines use the overhand cut and fill method to work upward from this level and the undercut and fill method to work downward this doubles the number of working faces in operation at any one time from a given level whether or not this is advisable depends on the ore rock strength wide ore bodies can also be mined using the underhand cut and fill method figure 6 5 17a and b the process is the same as described earlier but now cold joints occur between the individual drift floor pours generally whether it be engineered fills cemented fill or a paste fill it is better to avoid positioning drifts of the underlying layer directly under those above it is acceptable to shift the drifts sideways or drive them at an angle to those abov

e the latter case results in a basket weave pattern figure 6 5 17c undercut drifts are developed usually not on 90 angles but more commonly on 45 or 60 angles figure 6 5 18 in thick inclined or very wide ore bodies that are appropriate for overhand cut and fill mining vertical pillars are sometimes left to provide additional support between the hanging walls and footwalls in a method called postpillar mining figure 6 5 19 if possible the pillars are located in internal waste or low grade areas on the lowest slice an r p mine is created and the rooms are then filled a second slice is taken continuing along the vertical upward extension of the pillars and this level is then filled the process continues thusly this method is so named because the pillars appear as vertical posts surrounded by fill some consider it to be a subset of r p mining others to be a subset of cut and fill mining there are no set dimensions for the pillars and rooms in this mining method in figure 6 5 19 the small post pillars can vary from only 1 8 to 6 m 6 to 20 ft square and the rooms from 7 to 15 m 24 to 50 ft depending on the rockmass strength of the back the idea is to maximize the extraction ratio while maintaining a safe working area because of the presence of the surrounding fill even very tall and slender pillars can be quite strong the method is often used to maximize initial recovery from r p mining and therefore no additional recovery is expected by means of retreat mining or pillaring of course in all such cases measures must be taken to ensure that the pillars are correctly aligned and vertical sublevel caving mining method sublevel caving was used initially in the united states for extracting soft iron ores in the iron ranges of minnesota and michigan the method as practiced today differs significantly from that used earlier and should probably be given another name such as sublevel retreat stoping continuous underhand sublevel stoping or something similar that better reflects the process sublevels are created at intervals of 20 to 30 m 66 to 98 ft beginning at the top of the ore body and working downward on each sublevel a series of parallel drifts is driven at a center to center spacing on the same order as the level spacing from each sublevel drift vertical or near vertical blasthole fans are drilled upward to the immediately overlying sublevels the distance between fans the burden is on the order of 2 to 3 m 6 5 to 10 ft beginning typically at the hanging wall fans are blasted one by one against the frontlying material consisting of a mixture of ore from overlying slices as well as the waste making up the hanging walls and or footwalls extraction of ore from the blasted slice continues until the total dilution reaches a prescribed mineral cutoff level the next slice is then blasted and the process continued depending on ore body geometry the method can be used with transverse or longitudinal re

treats today sublevel caving is used in hard strong ore materials from which hanging wall rocks readily cave the key layout and design consideration is to achieve high recovery with an acceptable amount of dilution and ore loss the uncertainties of fragmentation and ore cavability present in panel caving discussed in the following section are removed since each ton of ore is drilled and blasted from the sublevels the method has been used most for mining magnetic iron ores which can be easily and inexpensively separated from waste however it has been and can be applied to a wide variety of other ore types sublevel caving layout ore is recovered by means of both drifting and stoping because the cost per ton for drifting is several times that for stoping it is desirable to maximize stoping and minimize drifting thus through the years sublevel heights have steadily increased until today stopes are up to 30 m 98 ft high whereas in early designs approximately 25 of the total volume was removed by drifting today in the largest scale sublevel cave designs that value has dropped to about 6 similarly sublevel intervals have increased from 9 to nearly 30 m 30 to 98 ft key to this development is the capability to drill longer straighter and larger diameter holes sublevel caving is an underhand method with all blastholes drilled upward the ore moves down to the extraction drilling drift by gravity a number of factors determine sublevel cave design sublevel drifts typically have dimensions width height of 5 4 m 16 13 ft 6 5 m 20 16 ft or 7 5 m 23 16 ft to accommodate lhd equipment in the example used to illustrate the layout principles the drift size is assumed to be 7 5 m 23 16 ft the largest possible blasthole diameter from the viewpoint of drilling capacity and explosive charging is normally chosen the maximum hole size in use today is 115 mm 4 5 in based largely on the capability to charge and retain explosive in the hole these large holes can be drilled using either in the hole or top hammer drills the large diameters and large drift sizes permit the use of tubular drill steel of relatively long lengths which minimizes the number of joints and maximizes joint stiffness so that the required long straight holes can be produced the largest ring designs incorporate holes with lengths of up to 50 m 164 ft a sample sublevel caving ring design is shown in figure 6 5 20 the sublevel drift interval is decided largely based on the capability to drill straight holes in this example it is assumed that the sublevel interval based on drilling accuracy is 25 m 82 ft when the sublevel interval has been decided it is necessary to position the sublevel drifts in this example drifts are placed so that the angle from the upper corner of the extraction drift to the bottom center of the drifts on the overlying sublevel is 70 this is about the minimum angle at which the material in

the ring moves to the drawpoint the resulting center to center spacing is 22 m 72 ft a single boom drill is assumed to drill all of the holes in the ring the inclination of the side holes is chosen as 55 although holes somewhat flatter than this can be drilled and charged the function of holes drilled flatter than 70 is largely 1 to crack the ore which is then removed from the sublevel below and 2 to reduce the maximum drill hole length holes flatter than 45 are difficult to charge because of the angle of repose of the ore at the extraction front in the final configuration of drill holes shown in the figure a buffer zone 1 m 3 3 ft wide has been left between the ends of the blastholes and the boundary to the overlying drifts and outer fan holes the layout is similar to that obtained using the theory of bulk flow described by kvapil 1982 1992 the fans can be drilled vertically or inclined from the horizontal at an angle typically 70 to 80 inclining the fans improves brow stability and access for charging the holes in the example the inclination of the fans is 80 and the burden is 3 m 10 ft to initiate mining of a new sublevel an opening slot must be made toward which the fans can be blasted several techniques can be used including blind hole boring and slashing fan drilling with increasing inclination angles until the production fan inclination is reached and creating an opening slot longitudinally along the hanging wall for transverse sublevel caving upon reaching the footwall the inclination of the fans is sometimes steepened to permit recovery of additional ore and to minimize waste extraction figure 6 5 20 shows the importance of good drilling precision if the forward or backward angular position due to incorrect initial alignment or in hole deviation exceeds 2 the ends of the longest holes find themselves in the wrong ring side to side angular deviations can mean that fragmentation is poor due to too little explosive concentration dead pressing of explosive and so on thus careful drilling is of the utmost importance for successful sublevel caving dead pressing of explosive means that the explosive density has increased by detonation of another nearby blasthole to the point that the explosive can no longer detonate this is most likely to happen with fairly insensitive explosives more sensitive explosives are more likely to detonate sympathetically rather than become dead pressed recovery and dilution sublevel caving lends itself to a very high degree of mechanization and automation each of the different unit operations of drifting production drilling blasting and extraction can be done largely without disturbance from the others specialized equipment and techniques have been developed leading to a near factory like mining environment as indicated earlier because every ton of ore is drilled and blasted there are no longer the uncertainties regarding cavability and degree of fr

agmentation as occur with block caving however a very narrow slice of blasted ore surrounded by a mixture of waste and ore must be extracted with high recovery and minimal dilution as is easy to visualize the ore at the top of the ring in the example is more than 40 m 130 ft away from the extraction point whereas the waste ore mixture lies only the distance of the burden in front of the drill ring pattern on the order of 3 m 10 ft with care recoveries on the order of 80 with dilution held below 25 can be achieved panel caving mining method the term panel caving here represents block caving suggesting the mining of individual blocks as well as panel caving used today mostly to indicate a laterally expanding extraction as one would expect the system has a great number of variants for panel caving the three most important elements of the extraction system are 1 the undercut level that removes the support from the overlying rock column 2 the funnel through which rock is transported downward to the extraction level and 3 the extraction level itself the basis for system design and performance is the degree of fragmentation present as the rock blocks enter the top of the funnel in the early days of block caving it was always considered essential that the materials be soft and cave readily today the trend is to use cave mining on ever harder and tougher ores as a result the engineer must thoroughly evaluate the ore body and tailor the design so that successful extraction will result this is the least expensive mining system as measured by extracted tons however it requires a large amount of testing during the feasibility study and thus only an engineer who is experienced in panel caving design should attempt to design a panel caving system extraction principles assuming the use of lhd equipment the major development on the extraction level consists of extraction drifts drawpoints and extraction troughs bells to simplify discussion it is assumed here that all drifts have the same cross section design is iterative and it is not always obvious where to begin in this case one begins with knowing or assuming the size of the material that must be handled the physical size of the loading equipment is related to the required scoop capacity which in turn is related to the size of the material to be handled if fragmentation is expected to be coarse a larger bucket and larger machine are required than if the extracted product is expected to be fine knowing the size of the machine one arrives at a drift size the orepass diameter should be 3 to 5 times the anticipated largest block size in order to avoid hang ups by this rule the extraction opening should be about 5 to 7 m 16 to 23 ft for block sizes up to 1 5 m 5 ft depending on density and shape such a block would weigh 5 to 10 t 5 5 to 11 st a large piece of equipment is required to handle such blocks it is typical for extraction drifts to be

sized according to width height ratios w h 4 3 5 4 and 6 5 for the machine in the example in this section the drift size would be about 5 4 m 16 13 ft or larger to begin the design of the extraction level one creates a grid of extraction drifts that are to be traversed by lhd units and the lines of the associated drawpoints in practice a series of circles of radius r corresponding to the draw radius of influence on the undercut level are first drawn figure 6 5 21 shows two such patterns for staggered coverage with the locations of the extraction drifts superimposed the value of r depends on the degree of fragmentation and is larger for large fragments and smaller for finer fragments this presents a design problem since fragmentation is generally larger in the initial stages than at the later stages of draw the degree of desired coverage is an important design factor the case of just touching figure 6 5 21a shows that triangles between the circles are not covered moving the circles to achieve greater overlap gives finally the case of total coverage figure 6 5 21b in this example it is assumed that r 7 5 m 25 ft and a square just touching drawpoint pattern is used figure 6 5 22 shows the locations of the extraction drifts drawpoint drifts and drawpoints on the extraction level for extraction drift 2 drawpoints 1 and 2 are associated with drawpoint drift 1 and drawpoints 3 and 4 are associated with drawpoint drift 2 the figure shows that two draw circles are associated with each drawpoint in this design for drawpoint 1 the draw circles are 1 and 4 for drawpoint 2 the draw circles are 2 and 3 in continuing the design example one must decide the orientation of the drawpoint drift with respect to the extraction drift figures 6 5 23 shows two possibilities involving the use of a 45 angle careful examination reveals that the choice affects both loading direction and ease with which openings can be driven a drawpoint entrance made at 60 to the axis of the extraction drift is at a very convenient angle from the loader operator s point of view some designs use 90 angles square pattern enabling loading to be done from either direction however although 90 pillars provide good corner stability loading is more difficult when considering different drawpoint design possibilities loading machine construction must be taken into account it is important for the two parts of the lhd unit to be aligned when loading to avoid high maintenance costs and low machine availability as mentioned previously the extraction level must be designed according to the degree of fragmentation expected so that the openings are large enough to permit extraction however large openings are prone to stability problems and because openings must last for the time required to extract the overlying column of ore it is important that design creation and reinforcement of the openings be done with care fortunately for h

arder rocks yielding coarser fragmentation although one expects the ore to be coarse the host rock is also strong and provides good construction material for softer rocks yielding finer fragmentation the openings should be smaller protecting the integrity of the openings is of the highest importance and is discussed in more detail in the next section there are a large number of design possibilities for the extraction level all involve the basic components of fragmentation radius of influence draw coverage machine size and drift size which should be examined in roughly that order undercutting and formation of the extraction trough in the undercutting process a slice of ore forming the lower portion of the extraction column is mined as this drilled and blasted material is removed a horizontal cavity forms beneath the overlying intact rock because of the presence of this free surface subhorizontal side stresses and the action of gravity the intact rock undergoes a complex process involving loosening crushing and caving the ease with which intact rock transforms into a mass of fragments is called its characteristic cavability one approach to addressing a material s cavability is to describe the size and shape of the area that must be undercut to promote caving the other and more important approach is to describe the fragment size distribution the latter is much more difficult to predict but ultimately more important from a design viewpoint in this section both the undercutting process and the design of the trough required to deliver the resulting fragments to the extraction level are described the simplest design is to combine undercutting and trough formation into a single step as described in the previous section a series of parallel extraction drifts are driven the center to center spacing of the drifts is determined by the size of the influence circles this example uses the plan layout of figure 6 5 22 the center to center spacing of the extraction drifts is 30 m 100 ft 4r a series of parallel trough drifts are now driven between the extraction drifts starting at the far end of the extraction block fans of holes are drilled and then blasted toward opening slots figure 6 5 24a shows an example in which the side angles of the fans are 52 and the resulting vertical distance between the extraction level and the top of the undercut is 20 m 66 ft the trough drifts and troughs can be created either before or after the extraction drifts are driven the latter case is called advance or pre undercutting an advantage of this design is that all development is done from one level an example of the use of this design has been presented by weiss 1981 most mining companies using panel caving have separate undercut and extraction levels figure 6 5 24b shows the undercut level designed as a rib pillar mine the rooms are 5 4 m 16 13 ft and the room center to center spacing is 15 m 50 ft in step 2 the

interlying pillars are drilled and blasted in step 3 extraction troughs are created to complete undercut trough development it is possible and often desirable to develop the undercut level first followed by the extraction level figure 6 5 24c shows an alternative design for the same basic extraction level layout a separate undercut level has been used with the undercut drifts spaced 30 m 100 ft center to center from these drifts fans of holes are drilled to form a trough the angle of the side holes has been chosen as 52 the undercut drifts are positioned directly above the underlying extraction drifts when the undercut has been created a sublevel caving type fan pattern is drilled from the trough drifts on the extraction level this completes development the total height of the undercutting in this case is 40 m 132 ft which has some advantages in the caving of harder rock types because it greatly increases the undercut volume thus inducing caving and the greater volume improves secondary breakage of the harder rock figures 6 5 25 and 6 5 26 show plan and section views of a more traditional undercutting and bell layout for panel caving the previous examples use primarily extraction troughs to demonstrate the principles involved a trough has the advantage of simplicity of construction but the disadvantage that additional rock is extracted during development this rock if left in place can provide extra stability to both the extraction drifts and the drawpoints drawbells are created rather than troughs the first step in drawbell construction is to drive a drawpoint drift connecting adjacent extraction drifts a raise is driven from this drift up to the undercut level fans of drill holes are then drilled from the drawpoint drift around the opening raise to form the bottom of the drawbell fans of holes are also drilled from the undercut drifts to complete bell formation a disadvantage of this method is that the amount of development and level of workmanship required is higher than for trough design as a result it is more difficult to automate for all designs it is important that a complete undercut be made if this is not done very high stresses can be transmitted from the extraction block to the extraction level causing major damage traditionally the extraction level has been prepared first followed by creation of the undercut and completion of the drawbells this procedure has a number of advantages unfortunately however very high near vertical stresses are created just ahead of the leading edge of the undercut these stresses are transmitted through the pillars to the extraction level and can induce heavy damage to the newly completed level the result is that repairs must be made before production can begin the concrete used for making the repairs is generally many times weaker that the rock that has been broken and the structural strength can never be completely restored an alternative to this proced

ure is to create the undercut first advance undercutting thereby cutting off the vertical stress the extraction level is then created under this stress umbrella where this has been done conditions on the extraction level are markedly improved over those where undercutting was done afterward both techniques have advantages and disadvantages but advance undercutting is preferred for most future mines block size the size of a block refers both to the height of the extracted column and its plan area in the early days of block caving block height was on the order of 30 to 50 m 98 to 164 ft with time this has increased to the point that extraction heights of several hundred meters are now being used or planned obviously because specific development is inversely proportional to the height of the block there are incentives to make the extraction units as high as possible within the restrictions imposed by factors such as ore body geometry and mineral type limits are also imposed by the life of the extraction points if the anticipated life of the extraction point is for example 100 000 t 110 000 st there is no point in selecting a block height yielding 200 000 t 220 000 st per drawpoint drawpoints can be rebuilt but it is best if they can last the life of the draw as mentioned earlier most caving today is done in the form of panel caving rather than the caving of individual blocks after the initial cave is started the lateral dimensions are expanded cavability affects the size of the undercut that must be created to get a sustainable cave relationships have been developed among rock mass characteristic hydraulic radius area perimeter and ease of caving it is possible unfortunately to begin initial caving and then for a stable arch to form the undercut area must then be expanded and or other techniques such as boundary weakening must be used to get the cave started again with a sufficiently large undercut area caving can be induced in any rock mass which is necessary but not sufficient for successful block caving another factor that affects the size of the undercut is the degree of fragmentation that results because panel caving is being considered for application to ever stronger rock types both of these factors cavability and degree of fragmentation must be satisfactorily addressed before any method is selected unfortunately the database upon which such decisions must be taken is very limited cave management cave management refers to maintaining control over how much is extracted from each drawpoint each day it involves a number of different factors the rate of draw is an important parameter in planning the required area under exploitation loosening of fragments appears to be a time dependent process and must be recognized in planning the draw the rate must not be so rapid that a large gap results between the top of the cave and the bottom of the block sudden collapse of the rock above can result i

n disastrous air blasts in high stress fields too rapid draw can cause rock bursting conditions in the drawpoint area there is a zone in which the height of the column under draw increases from near zero where extraction is just beginning to the full column height this is followed by a zone in which the ore column height decreases to near zero where extraction is complete it is important to maintain the proper ratio of draw height to slope distance to avoid early introduction of waste from above poor cave management can also mean buildup of high loads in various areas and subsequent stability problems typical rates of draw taken from the available literature are on the order of 0 3 to 0 6 m d 1 to 2 ft d the proper sequencing of undercut and extraction is a very important aspect of cave management unfortunately relevant design guidelines are difficult to obtain from the literature an important design consideration for the extraction level is the means by which oversize is handled there are a number of different problems to address the first concern is the management of true hang ups at the extraction points sometimes these can simply be blasted down by careful placement of explosives at other times the boulders must be drilled first this is not a simple procedure and involves danger to both machines and humans the second concern is where and how to handle the movable oversize these blocks can be 1 handled at the extraction points 2 moved to a special gallery for blasting 3 moved to an orepass equipped with a grizzly and handled there or 4 dumped directly into an orepass for later handling all variations are used and each company has its own procedures initially the sizes of blocks arriving at the drawpoints are defined by natural jointing bedding and other weakness planes as blocks separate from the parent rock mass they displace and rotate with the loose volume occupying a larger volume than does the intact mass the swell volume is extracted from the extraction points providing expansion room for the overlying intact rock loosening eventually encompasses the entire column as the column is withdrawn the individual blocks abrade and split resulting in finer fragmentation than in the early part of the draw the initial fragmentation due to initial fractures in the rock is called primary fragmentation as the column moves downward and new breakage occurs the resulting fragmentation is called secondary fragmentation data concerning this transition from primary to secondary fragmentation are very difficult to obtain each mining operation is unique in some aspect ownership ore body location size geometry grade mineralogy hydrology or geotechnical parameters or constraints such as environmental regulations however all mining operations are alike in one way the prime objective of any mining project is to maximize the return on investment this is the golden rule or in the language of mining

the investor s law of conservation a mineral resource by definition does not become a mineable reserve unless the mineral or minerals can be extracted economically and legally the choice of the most economic method is dictated primarily by the type of deposit to be mined the mining rate and the physical characteristics of the deposit a comparison of surface methods must take into consideration two prime objectives of mine planning to develop the most economic plan for the overall project that will maximize the return on the monies invested and to maximize the recovery of the resource the purpose of this chapter is to discuss the relative advantages and applications of one method versus another as well as some of the more important evaluation parameters that must be considered in any method comparison factors and conditions open pit mining and strip mining are the two most dominant types of surface mining methods in the world accounting for approximately 90 of the surface mineral tonnage strip open cast mining is used for large tabular flat lying ore bodies or mineral seams such as coal that are relatively close to the surface backfilling of these mines is usually economically feasible and desirable as part of the concurrent reclamation requirements open pit mining is typically applied to disseminated ore bodies or steeply dipping veins or seams where the mining advance is toward increasing depths backfilling usually cannot take place until the pit is completed even then the prohibitive cost of filling these pits with all of the waste rock removed at the end of the mine life would seriously jeopardize the project s economics very few large open pits in the world could support such a costly handicap aggregate mines produce a bulk commodity for local markets these quarries employ the same mining methods as other open pit mines seam deposit mining like coal uses different planning methods and terminology as compared to open pit mining for base and precious metals however all surface mining methods have one common element mining proceeds until the economic or legal limits are reached the method of determining the economic limit is different for the two types of deposits coal for example in most instances is readily distinguished from waste rock and is of a near uniform quality and value per ton mined in each seam therefore mine design limits are determined by the break even strip ratio commonly defined as cubic yards cubic meters of waste to tons metric tons of coal that can be mined since 1 ton of coal is approximately equal to 1 yd3 1 t 0 84 m3 the ratio is similar to a volume ratio the pit limits for a disseminated deposit such as base or precious metals is determined by a break even economic analysis of the value of a finite incremental pit expansion this economic analysis is a two step process where first the net operating value is determined for each ton of material in a block model a

bove and below the ore cutoff grade second the pit limits are determined by expanding the pit shape in increments to a break even point where the net operating margin equals the cost of waste stripping given that each ton of ore has a different grade and net value a break even stripping ratio cannot be used in the determination of the pit limits as a final test the increment between an existing pit wall to the economic limit must have sufficient width for the proposed mining method complicating this matter further is the potential for multiple ore processes such as milling and leaching of the same ore type milling is a higher cost process method used to increase metal recoveries whereas leaching is a slower process with correspondingly lesser recoveries and lower costs therefore the mining engineer must have sufficient knowledge of the overall project from ore geometries to the process extraction methods in order to maximize the project return on investment this includes an awareness of all cost and recovery parameters as well as the factors that affect the value of the salable product for example where process capital and operating costs are high ore selectivity and the choice of appropriate mining method play important roles in mine design mining cost increases that may be required to optimize the project s return can be more than offset by reduced process capital and operating costs per unit of salable product the total amount and type of material to be moved per time period coupled with ore body geometries are the prime criteria in selecting the most economic and applicable mining method other factors that affect the type of mining method as well as type and size of equipment include topography remoteness of operation support infrastructure availability and cost of skilled labor climate and altitude environmental legal or physical boundary restrictions groundwater quantities slope stability and ground conditions investment risk blasting restrictions life of operation and waste rock disposal topography can play a major role in selecting the most appropriate mining method generally the more rugged the topography the more limited are the choices of mining method and level of overall production remote operations require substantial infrastructure to maintain the standard of living needed to attract a wellqualified work force therefore the mining and processing methods selected must result in high productivities per employee shift better ore selectivity in the mine will reduce process plant requirements and could therefore reduce overall staffing requirements if the ore grade can be raised sufficiently the ore may be transported off site for processing at a less remote location large equipment can be used for the bulk of the waste stripping and small equipment employed to selectively mine the ore where labor costs are high the same philosophy can be applied lower labor co

sts and the level of labor skills available can also affect the choice of mining method for example front end loaders and mechanical drive trucks will require less diverse maintenance skills as compared to the labor force needed to maintain electric hydraulic shovels and electric haul trucks with adverse climatic conditions mining operations may be seasonal seasonal and short life operations may use contractors for all mining operations or for only the bulk waste stripping environmental and social considerations play an increasingly important role in the type of mining operation permissible open pits in populous areas may require backfilling whereas the same type of operation in a remote location may have minimal environmental constraints for example hydraulic mining outlawed in california united states in 1884 was used to remove the unconsolidated overburden at the bougainville copper mine papua new guinea in 1969 1971 environmental and aesthetic conditions may be so restrictive in some areas as to eliminate surface mining options altogether in some cases this forces the use of underground mining which will result in reduced mineral resource recovery groundwater conditions and slope stability can play an important role in mining method options mine design production scheduling and equipment selection a large dragline may not operate safely where ground conditions are wet and unstable and the risk of slope failure is high if large water inflows are anticipated mining may have to be accelerated to avoid high pumping costs from a lengthy operation ore can be stockpiled at the processing plant and treated months or even years after the mine is completed the risk of investment varies from country to country however if risks are high then capital expenditures are usually minimized front end loaders with shorter useful lives and lesser capital costs are preferred over large shovels and draglines the larger the project and the longer the development period the greater the investment risk blasting restrictions due to the operation s proximity to populous areas or in foreign countries with explosive use constraints can also affect mining method selection continuous mining machinery similar to pavement cutters or ripping can be used in some circumstances as an alternative to drilling and blasting available waste disposal areas may be located a significant trucking distance from the surface mining operation a viable alternative in this situation is crushing and transporting of the waste via a line of conveyor belts and this in turn can affect the mining methods where the ore has a high value per ton in excess of 100 times the cost of mining very high strip ratios can be economically justified as well as underground mining methods the surface mining method in these circumstances is designed to be primarily a waste removal operation ore mining is a small cost component of the overall operation and great emp