Coal Resource Recoverability
A Methodology
U.S. Bureau of Mines Circular 9368
PROGRAM DESIGN
After examination of the USGS-KGS Coal Availability project in early 1989, a preliminary recoverability methodology was developed, as were map needs and formats, preliminary work schedules, and mining restriction criteria. Geological information was collected from the KGS, in Lexington, KY, and the locations of all mining, wash plant, and waste-disposal sites in the Matewan and surrounding quadrangles were obtained from the regional office of the Kentucky Department of Surface Mine Reclamation Enforcement (DSMRE), in Pikeville, KY. The Matewan quadrangle was toured so that (1) the locations of wash plants, railroads, highways, haulage roads, streams, and rivers could be verified, (2) trends in the terrain and haulage routes could be identified, and (3) any other pertinent mining infrastructure could be noted. Several mines in and around the Matewan quadrangle were evaluated for operational details. An in-depth literature and industry search for state-of-the-art mining methods and equipment and current production rates was undertaken, and recovery rates of similar mining configurations in other localities were investigated. Normal planning variables, such as seam thickness, floor and roof quality, seam hardness, seam quality, parting and coal content within a seam, and depth of seams, were collected. Industry and government sources were used for equipment costs and cost indexes, productivity, transportation costs, coal quality versus coal realization, taxation, etc.
The data from all the various sources were then evaluated in connection with the USGS methodology of determining coal resource availability. This methodology was then modified and expanded, the results of which include a different definition of coal resource availability, and herein lies a major difference between the USGS program and Bureau program methodologies. This difference between the USGS Coal Availability program and the Bureau Coal Recoverability program is the Bureau's use of total coal seam thickness in determining mining methods, resource recovery, and applicable mining costs. The Bureau assumptions and methodology reflect recovery of some coal in the mining of full seams that would have been excluded in a coal-only scenario (i.e., seam parting of up to 50% of the total seam thickness is included in the minable resource number). Also, the Bureau did not limit the mining of the coal if the coal had high sulfur. or high ash content; instead, operating costs became the limiting factor for the reserve. Thus, high sulfur coal is limited in reserve by its lower sales price. Ash content, denoting inseam parting and out-of-seam dilution, is accounted for by (1) the additional cost of washing the run-of-mine (ROM) product, and (2) the percentage of ROM product washed out during cleaning, thereby effectively lowering the clean or salable tonnage recovered and raising the cost per ton to produce the coal.
As a result of the redefinition of coal resource availability, the Bureau estimation of the total original in-place, resource for the Matewan quadrangle is 21% more than the USGS estimate; however, due to significant differences in estimated resource losses determined for previously mined resources and restrictions to future mining, the Bureau estimate of the remaining minable resource is less than the USGS estimate by 14%. In any event, this revised minable tonnage, derived from USGS and KGS coal availability data, formed the foundation upon which the Bureau's recoverability program fully developed.
The Bureau recoverability methodology, which consists of removing the resources lost during mining and washing, results in recoverable resources. Mining economics are then applied to the recoverable resources to determine the amount of coal that can be mined at a profit. To determine the resource recovery, mining assumptions were made and preliminary mine plans were developed for each seam. Coal mining and washing recovery models were developed and applied to the resource data. The recoverable resources were then incorporated into the mine cost models to calculate the economically recoverable resources (i.e., reserves) at different levels of the coal market. This methodology easily acknowledges the fact that the reserve base, or "economically recoverable coal base," will change with changes in energy prices.
Site and mine data from field investigations provided the base information necessary to evaluate the effect of the land use and technical restrictions on the mining conditions and recovery rates. Observations by Bureau investigators, the USGS, and the KGS prompted the authors to add and alter restrictions used in the evaluation.
The USGS identified several land use restrictions that the Federal Surface Mining Control and Reclamation Act of 1977 (SMCRA) and the Kentucky Surface Mining Law of 1982 would apply to coal resources. Rather than divide the restrictions by surface or underground mining, this report methodology divides those restrictions by industrial and environmental plus social considerations. Although some definitions were modified, the principal difference between the Coal Availability and the Coal Recoverability studies lies in the definition of the coal seam itself. For years, the Nation's coal resources have been defined bycoal-only thicknesses within the seam (13, 17). The question is stated then as follows: If presented with varying amounts of noncoal lithology (shale, siltstone, etc.) within a seam, at what ratio of noncoal material to coal does one consider the seam no longer minable?
Field investigations of mines in the central Appalachian region that wash coal prior to sale, particularly in the Matewan quadrangle area, found that 50% to 65% recovery of the ROM coal seam was fairly common. For that reason, the assumption was made that if a coal seam contained 50% or more coal, by thickness, it could be mined. Since this is an economic restriction, it will be addressed where wash plant recoveries are factored into the cost to produce a ton of coal from a particular seam.
The industrial restrictions encountered in the Matewan quadrangle are powerlines, pipelines, oil and gas wells, and highways and railroads. Although these restrictions can be abated through relocation or positive abandonment (well plugs), the high cost normally associated with abatement usually prohibits economic recovery of the resource. The KGS and USGS staffs digitized the locations of major powerlines, pipelines, and oil and gas wells. No major highways or railroads were digitized since none would be affected by mining. Therefore, the Bureau's industrial restrictions are identical to those of the USGS.
Environmental and Social Restrictions
Environmental constraints addressed in the study of the Matewan quadrangle are streams and premining ground slopes where surface mining could take place.
For streams having a mean annual flow of 5 cfs or greater, mining is not allowed within 100 ft of the highwater level. All surface resources inside the buffer zone were considered restricted. As with many industrial constraints, variances can be obtained to divert or reroute streams so that surface mining can take place. However, construction and subsequent reclamation are generally cost prohibitive.
Present surface mining laws mandate the replacement of overburden to the approximate original contour, unless a variance is obtained (such as in mountain-top removal stripping). The maximum slope for reclamation is dependent upon the composition of the rock types in the overburden above the coal seam and interburden between coal seams (if more than one seam is mined) and the amount of water present. Stable reclaimed slopes are normally 62% to 70% (rise to run). When accurate digital terrain models are available with the outcrop of the coal seam, slope measurements can be done by computer. Ground slope measurements were conducted perpendicular to the contour across all outcrops, using the USGS 7-1/2-minute Matewan topographic map. No average slope measurements exceeded 62% (32) in minable areas; therefore, no slope restrictions were placed on the minable resources.
Social constraints are cemeteries, towns, and residences. Variances can usually be obtained to move the restriction to another location, but only at significant cost and time. These mining restrictions were accepted in total as presented by Carter and Gardner (16). Table 1 shows the restrictions and mining assumptions used in the USGS and Bureau studies.
Technical restrictions are of two categories: coal seams that are too thin to mine with today's equipment, and restrictions where the roof or floor would be unstable due to past or future mining in seams adjacent to the one under consideration. In the Matewan quadrangle, the depth to the resource and geological factors were also considered.
This restriction was changed significantly from that reported by Carter and Gardener. They used the accepted mining thickness parameters-14 in of coal for surface mines and 28 in of coal for deep mines-for determining coal resources (15). During the authors' evaluation of central Appalachia properties, deep mines were observed to be operating in a 24-in seam of coal, and strip mine operators used a seam thickness cut-off of 12 in (usually in multiple-seam stripping). In all cases, the obvious restricting factor is economics. Further, since the USGS's definition of coal resources applied to coal only, no parting material greater than 3/8 in thick was included in the 14-in and 28-in thickness restrictions. Fortunately, the parting thicknesses had been recorded in the KGS data base, and the USGS was able to add those partings to the coal to obtain a real seam thickness model. The authors used a 12-in minimum seam thickness for surface mine planning and a 24-in minimum seam thickness for deep mining. As noted earlier, a cutoff of 50% parting in the seam was assumed as a mine planning parameter. The addition of the partings to the seam contributed to the large variance between Bureau resource calculations and the calculations of the USGS.
Several stability factors were addressed where the rock strength was known
to be too low to support mining close to vertically adjacent seams or previous mining. The
KGS felt that a 40-ft thickness of interburden between seams was a good average where
knowledge of specific rock strengths between the seams was lacking.
In several mine evaluations, it was found that some mines operated within 25 ft or less of
overlying abandoned mines; however, even with good pillar alignment from seam to seam,
poor roof conditions occurred in some interacting mines where interburden thicknesses were
greater than 60 ft. In some cases operators were attempting to mine over abandoned mines
with less than 60 ft of interburden. More often than not, floor and roof failure were
major problems. The authors considered the 40-ft interburden thickness as a minimum and,
further, restricted the mining sequence to be from the top of the stratigraphic sequence
downward.
Perennial drainages were considered to negatively affect the overburden through deep
weathering, which contributed to ground control problems. Several potential problem areas
were identified, and buffer zones of 200 ft were decided upon and designed as appropriate
restrictions.
Some past coal resource definitions assumed that deep mining was limited to depths of less
than 1,000 ft. Although many coal mines are operating at depths in excess of 2,000 ft, the
original resources were calculated only to a depth of 1,000 ft. In the Matewan quadrangle
no coal resources below the lowest drainage (Tug Fork of the Big Sandy River) were
considered by the USGS for evaluation because of this restriction. Subsequent quadrangle
evaluations have not used the 1,000-ft, nor the below-drainage, parameters as
restrictions.
In addition to depth restrictions on deep mining, the USGS used a depth restriction of 200
ft for surface minable seams to expedite the resource calculation. However, due to
economics, a highwall ratio was found to define more realistically the surface minable
resource. The authors found that by comparing effective stripping ratios (bank cubic yards
of overburden moved to tons of coal recovered) to the highwall ratio (thickness of
overburden to thickness of coal seam), a 20:1 highwall ratio was consistent with present
economic stripping limitations. On the other hand, the 200-ft depth limitation could
produce highwall ratios of as great as 200:1-well outside the economics of mining.
The final stability category reviewed was that of geologic factors such as faulting,
steeply dipping seams, splits in the coal seam, etc. These factors were not major problems
in the Matewan quadrangle and were not considered as a restriction to mining.
Two types of barrier restrictions were defmed. The first restriction is the 50-ft mandatory safety barrier around old mines to prevent unexpected water inflow and interruptions in mine ventilation. The second was calculated from the feasibility-type mine plans; it included barriers left between production panels, barriers left to protect ventilation, haulage, and travelways, and barriers left between the underground mine and the highwalls of the adjacent surface mine. This second barrier restriction was not addressed by the USGS Coal Availability program and amounts to approximately 7% of the total original resource in the Matewan quadrangle. The restrictions are summarized in table 2 and detailed in table 3.
Local and practical mining methods were employed in the feasibility-type mine planning. Four surface mining methods were examined: area stripping, mountain top removal, contour stripping, and auger mining. Contour stripping, auger mining and mountain top removal were considered applicable, area stripping was not. Continuous miner and longwall operations were planned for deep mining, due to their high productivity, but conventional mining was not considered; however, it is recognized that conventional methods (drill, cut, and shoot) would have to be employed where geological conditions dictated. Recovery rates were determined for the applicable mining methods, and for the deep mining methods simple regression curves were fitted to the accumulated data for production rates these were used to assign production rates for the deep mining methods by configuration and mining height. Surface transportation of coal, both washed and ROM, was addressed by measuring haulage distances to known wash plant locations and rail loadouts.
For many years a convenient rule of thumb of 50% has been used when estimating the amount of recoverable reserves in a loosely defined resource or mining area. This type of average does not consider either the minability of deposits, changes in technology, or changes in demand for coal. It is impossible to predict accurately the eventual recoverability of all coal deposits; however, a current average of recoverabilities depicting today's technology and economics will be useful for estimating deposits that are presently minable.
Lowrie (18) makes a very good observation that bears repeating:
The rationale, traditionally, for using the convenient figure of 50 percent as an acceptable approximation of average recovery is that it compensated for losses which were not ordinarily included at individual mines. This is an indefinite generation which allows broad interpretation and perhaps erroneous exclusion of potentially recoverable deposits. Marginal deposits or those not ordinarily considered profitable or minable today, such as those behind prior workings, of poor quality, under adverse physical conditions, undermined, or underlying surface structures and so on, may be minable later rather than "lost.' Subsequent reevaluation of deposits as a result of changes in technology and economics should be made from a category other than "lost', which has a permanent connotation. Definitive evaluation of losses attendant in mining such deposits should be deferred until the time of exploitation.
Each mining method considered for this project demonstrates a recovery unique to that particular method. For this reason, variations of each method were investigated. The investigation included creating several "typical" models and calculating the recoveries, analyzing published and unpublished recovery data from government and industry sources, and drawing upon personal experience.
The mining methods considered in this report are:
1. Contour strip (surface), with two coal thickness categories: 12 in to 36 in and greater than 36 in.
2. Auger (surface), with two coal thickness categories: 12 in to 36 in and greater than 36 in.
3. Continuous minor, five-entry system (underground), using two different pillar sizes: 40 ft by 40 ft and 80 ft by 120 ft.
4. Longwall, three entry development (underground), using two different pillar sizes: 40 ft by 40 ft and actual mine plan dimensions.
The following definitions that pertain to coal recovery, as used for this study, are derived from the USGS (17):
1. Recoverable coal-The coal that is or can be extracted from a coal bed during mining. The term "recoverable' should be used in combination with "resources" and not with "reserves."
2. Recovery percent-The percentage of coal extracted from a bed where the total tonnage originally in the bed is equal to 100%.
3. Recovery factor-The estimated or actual percentage of coal, expressed as a decimal, that can be or was extracted from the coal originally in a bed or beds of an area, mine, district, field, basin, region, province, township, quadrangle, county, State, political province, Nation, and/or the world.
The following 6 factors are identified as significantly affecting recovery percentages in underground mines (18):
1. Pillaring system.
2. Top rock and conditions.
3. Bottom rock and conditions.
4. Marketability.
5. Coalbed thickness.
6. Productivity.
The following 13 factors are identified as significantly affecting recovery percentages in strip mines (19):
1. Top coal removed with overburden.
2. Bottom coal left in place.
3. Coal lost in ribs between cuts.
4. Losses due to partings.
5. Losses due to other factors such as equipment limitations and spillage.
6. Outcrops (coal exposed to weathering's degradation).
7. Utility, railway, and highway rights-of-way.
8. Streams.
9. Underground workings.
10. Oil and gas wells.
11. Property lines.
12. Buildings.
13. Other, such as landslide areas and cemeteries
The following 5 factors are identified as significantly affecting recovery percentages in auger operations (19):
1. Auger diameter.
2. Auger spacing.
3. Average auger penetration.
4. Coalbed thickness.
5. Length of coalbed available for augering.
A "typical" mine plan was drafted to scale with sufficient detail to
incorporate the standard pillars, entries, and development works necessary to provide a
representative scenario of the particular mining type under study. The mine plan was
digitized, using the GSMAP computer program (20), to provide two dimensional digital
information of the mine plan. The digitized data was reproduced on paper by a pen plotter
to verify its accuracy. The areas of all pillars were then calculated by the GSMAP
program. The accuracy and methodology of digitizing using GSMAP were checked on three
seams using multiple-pass planimetering. The mining and restriction areas were nearly
identical between methods.
Standard, or "average,' recovery factors were established for each of several
conditions and mining methods. It is recognized these are not, and cannot be considered,
definitive under all conditions, but they do reflect today's technology and economic
conditions more accurately than the often used 50% factor.
The following mine recovery factors are used for this study:
1. For contour strip mines-78% where the coal seam is less than 36 in thick, and 93% where the coal seam is greater than 36 in thick.
2. For auger mines-30%.
3. For continuous miner mines-62% where the pillars are 40 ft by 40 ft, and 57% where the pillars are 80 ft by 120 ft.
4. For longwall mines-84% where the pillars are 40 ft by 40 ft, and 78% where the pillars vary in size.
The resource recoverability calculations consist of the entire process of determining, from raw data, the amount and salability of coal from the area in question. The viability and accuracy of the calculations and associated modeling is dependent upon (1) the existence of a proven geologic model of the coal bed, (2) mining methods, (3) mining recovery, (4) mining sequence, (5) coal quality, and (6) dilution factors. Also influencing the model are coal washability parameters, wash plant recovery, mine size optimization, equipment configurations, land ownership, and all of the various environmental, social, and technical restrictions to mining. Equally important is the care with which the data are collected and used. Throughout the modeling process, utmost care was taken to assure that current, state-of-the-art methods were employed and that error was minimized.
The rate of production for surface and underground mining operations has long been discussed, studied, and computerized. Information on many of the factors influencing production within a specific coal seam, as shown in table 4, were not available during the data collection. Rather than calculate idealized production rates from the influencing factors, actual ROM production rates from Bureau evaluations and from industry and government publications were used. Contour strip and auger mining production rates were determined from mine evaluations and heavy equipment production charts and were fairly constant. Production rates for deep mining methods employing continuous miners or longwaus varied over a wide range. Longwall productivity appeared to correlate with mining height, and linear regression analysis proved this assumption to be correct, with very little scatter of the data points from the mean. Continuous miner productivity was found to have fairly good correlation to mining height after other mining factors had been accounted for. These other factors (e.g., available spare or standby equipment, supersections, remote controlled equipment, extremely good mining conditions, and retreat pillar robbing) foster significantly higher production rates than normal. Development mining, development for longwalls, and some captive mines, have lower than normal production rates. When the mine circumstances were addressed, the continuous mining productivity could be broken out into high, normal, and low production rates, and linear regression analysis suggested good correlation between mining height and production rate. These regression analyses are shown in figures 3(A) through 3(D).
Average production rates were determined for several ranges of mining heights and mining methods. Tables 5 and 6 show the production rates used in this study. These seam height configurations were used to specify production equipment sizing and capital costs for the mine costing models.
A feasibility program utilizing the three main Upper Elkhorn seams was begun in 1989.
Isopach maps of the coal, interburden, overburden, previously mined areas, and land use
restrictions were obtained from the USGS. Generalized mine plans were developed and the
close proximity of the three seams to each other allowed the interaction from seam to seam
to be studied during mine planning. This feasibility study demonstrated that preliminary
mine plans and restrictions could be merged and an estimate of the recoverable resources
could be made. Based on these findings, the program was expanded to include the remaining
minable seams in the Matewan quadrangle and to develop mining cost models that would allow
estimates of the economically recoverable resources.
The surface mining configuration was planned to operate with two separate fleets of
equipment in two different locations at the same time. This procedure utilized support
equipment, enhanced coal production and blending, and allowed sequencing of auger mining.
The planning process assumed common ownership of surface and mineral rights and assumed
that mining would be done in a logical, orderly manner for maximum recovery of the
resource. Only contour stripping methods were applied in the Matewan quadrangle. Neither
area mining nor mountain top removal was planned. The production parameters in contour
stripping were dependent upon seam thickness. Seams 36 in thick or greater were assumed to
have stripping rates of 6,500 BCY per shift, whereas seams that ranged in thickness from
12 in to 36 in had stripping rates of 5,200 BCY per shift. Mine bench widths varied, but a
minimum of 80 ft was used for all stripping. The 80-ft minimum width allowed endloaders to
safely load trucks and provided the trucks with enough room to turn around on the bench.
Since the average ground slope is about 26' (48.8%), the minimum bench width could be
maintained with seam thicknesses of less than 24 in. These conditions, along with the thin
- interburden between many of the seams, would allow multiple seam stripping. For this
reason, the 80-ft minimum width bench was used for all strip-minable resources with seam
thicknesses between 12 in and 24 in. During the surface planning, logical portal locations
were noted for underground development. Surface mine production was evaluated using a work
schedule of 8 hours per shift, 2 shifts per day, 5 days per week (240 days per year). It
was assumed that the surface mining equipment would develop the "face-up" area
for the deep mine portals, mine offices, and applicable stockpile areas. Reclamation plans
would be developed that would leave enough bench area for production haul roads from the
deep mine stockpiles.
Auger mining followed closely behind the stripping operations and utilized the coal
production equipment and support equipment. It was assumed that the stripping operation
left the highwall and bench in a condition that would require little maintenance by the
auger operations. Twin-headed augers were the chief production equipment, having a
compliment of drill heads and auger steel matched to the coal seam thickness. Average
recovery rates of 30% and average depth of holes of 60 ft were assumed. Highwall miners,
while not specified, could be an option for the thicker seams.
Deep mine development was laid out with five mains, 18 ft to 20 ft wide, having normal
offset barrier pillars and production panels employing room-and-pillar mining methods laid
out at right angles to the mains. Submains were laid out beneath and parallel to the
mountain ridges (spurs). Minimum panel widths were five entries; as the panels approached
the outcrop, the minimum advance was with two entries. Only mining with continuous miners
and longw@ was considered. The use of conventional roomand-pillar mining (drill, cut, and
shoot) is fairly common in the Matewan quadrangle; however, wherever possible, operators
are turning to continuous miners. Also, it is recognized that some coal seams are best
mined by conventional methods due to particular geological conditions. That information
was not available to the authors in the detail necessary for planning; therefore,
conventional mining methods were not addressed in this study. The optimum-size operation
was assumed to be two continuous miner production units for continuous mining methods, and
two continuous miners and one longwafl panel for a longwall operation. Minimum mining
heights for continuous miners was set at 24 in of seam thickness and 5 in of out-of-seam
dilution. The longwar minimum seam thickness was 42 in, with 3 in of out-of-seam dilution.
Because of mining height restrictions on equipment, continuous miner operations had four
separate height ranges: 24 to 42 in, 42 to 72 in, 72 to 96 in, and greater than 96 in.
Also because of restrictions on equipment, longwall mining had three mining heights: 42 to
72 in, 72 to 96 in, and greater that 96 in. Equipment specifications, capitation,
production, etc., were all determined by these mining heights in the recovery and costing
models. In addition to the minimum thickness of 42 in of seam height for longwalls, at
present a machine size limitation (motor to horsepower limitations), panels were laid out
with three entries on both sides of the panel, a minimum width of 550 ft, and a minimum
length of 4,000 ft. Areas considered for longwah mining had to contain enough coal
resource to depreciate the longwar equipment used in that seam. AR deep mine production
was scheduled for 8 hours per day, 2 shifts per day, 5 days per week (240 days per year).
Production rates are noted in tables 5 and 6
by mining method and height.
Development methodology for the basic mine plans. is listed below:
1. Understanding of the coalbed to be planned and the supe@acent and subjacent minable seams.
2. Identification of the effect of the land use restrictions on the subject bed.
3. Ground slope determination, along the outcrop, to outline areas having an undisturbed slope greater than 32'-for technical restrictions and for determining the highwall stripping ratio cutoff.
4. Surface mine design using the maximum 20:1 highwall ratio to determine the bench widths and stripping areas. Where the highwall ratio could be greater than 20:1, a minimum bench width of 80 ft was used.
5. Auger mining planned to coordinate with the stripping. Production tonnages were calculated by bench length, seam thickness, depth of auger holes, and normal recovery rate.
6. Underground mine development planned using working site areas prepared during stripping. Mains and submains were planned to be driven parallel to and beneath the mountain ridges.
7. Evaluation of seams for areas that met or exceeded the minimum longwall criteria. When a potential longwaff area was found, the panels were laid out and the continuous miner plans were designed around the longwaff plan.
8. Integration of the coal-haulage zone map with the mine plans. Original resources, restrictions, and minable resources were calculated for each haulage zone.
Using the above guidelines, the first mine plan maps were constructed for the Upper Elkhorn 1B seam (fig. 4). Planning for the Upper Elkhorn 2B seam followed, and conflicts between the mining plans for the 1B and 2B seams were resolved where necessary (i.e., where the interburden between the two seams was less than 40 ft). Next, the Upper Elkhorn 3.5 seam was planned. Since the interburden between the 3.5 and 3D seams exceeded 40 ft in thickness, there were no mining plan conflicts between these seams; however, the interburden between the Upper Elkhorn 3.5 and Williamson seams was thinner than 40 ft in many areas and caused major mining plan conflicts between the two seams. In this same way, mine plan maps of the other eighteen coal seams were subsequently constructed.
Topography in the Matewan quadrangle is rugged, having valley wall slopes ranging from 1Y to 320 (27% to 639o) and valley floors sloping from r to 13" (4% to 23%). Four major ridges control the watersheds in the quadrangle. During the initial tour and investigation of the quadrangle, the authors noted that coal was not transported over these main ridges. Indeed, only one ridge was crossed by a paved highway. Therefore, for the sake of efficiency, local practice was for all loaded trucks to haul coal downhill by the most convenient route to their destinations. Locations of preparation plants and train loadouts were obtained from the DSMRE permit maps. State, Federal, and county roads were located on the USGS 7-1/2-minute topographic map. After reviewing the geology and the potential for mining within the different watersheds, the quadrangle was divided into 16 haulage zones (fig. 5). The geographical center of all minable coal beds was located in each haulage zone and assumed to be the mine coal stockpiling area. Haulage distances were estimated from the stockpile area to the closest preparation plant and/or rail loadout. In this haulage scenario, the following assumptions were made
1. The closest wash plant always had the capacity to handle the delivered ROM coal. 2. Only wash plants with rail loadouts were used, thus eliminating double handling of the coal. Studies were conducted to see if the costs would be less to haul to a nearby mine washing facility, clean the coal reload the trucks, and then haul to a rail facility. However, none of the Matewanquadrangle coal seams contained enough ash to justify loading the coal in a truck the second time. Of course, this justification is valid only if enough washing capacity is available from facilities with rail loading capabilities.
3. No new washing facilities were necessary; however, it was assumed that all wash plants were operated at high efficiency rates, with idealized labor needs, and at normal operating costs.
4. All surface haulage of coal was done by contractor owned- and-operated trucks. Coal haulage contract costs for the southeast Kentucky, southern West Virginia, and southwestern Virginia areas were calculated as a fixed cost per truck plus a cost per ton-mile. These are included in the mine cost models as a basis for calculating the coal haulage costs. Escalating factors are used in the models to keep costs on a current level. Table 7 shows the oneway haulage distance and the cost assigned to each zone, for the Matewan quadrangle, as used in this study.
Information concerning the quality of the coals within the Matewan
quadrangle ranges from extensive to nonexistent. Due to the variabifity of coal
characteristics throughout central Appalachia, an attempt was made to determine coal
quality for the various seams from analyses specifically from the Matewan quadrangle. When
analyses from within the quadrangle could not be found, however, coal quality analyses
from adjacent or nearby areas were then used. As may be imagined, the most productive coal
seams have the greatest available amount of coal quality information, whereas the
nonproductive seams have little or no coal quality information specific to the quadrangle.
Current information indicates that all the coals within the quadrangle are low- or
medium-sulfur, high-volatile, bituminous varieties. If the standard of 1.2 lb of S02 per
MBtu is considered, there are almost equal amounts of compliance coals and noncompliance
coals within the quadrangle. Since coal quality data for all but two of the seams is
limited, the information represents a restricted characterization of the coal within the
Matewan quadrangle and is not statistically supported for precise quality portrayal.
Raw data accumulated by the KGS and the USGS are in the NCRDS. The data are retrieved as ASCII (American Standard Code for Information Interchange) data files, as needed for input to the system's GARNET (Graphic Analysis of Resources using Numerical Evaluation Techniques) computer program, and transmitted to the computers at the Bureau Intermountain Field Office in Denver, via the USGS computer linking networkGeonet. The following data are provided:
1. Coal seam extent (outcrop) and thickness (isopachs).
2. In-seam parting distribution and isopachs.
4. Overburden and interburden isopachs and structure
contours on the bottom of the coal seam.
5. Mined-out areas and their barrier piflars, both surface and underground.
6. USGS-defmed environmental and land use mining restrictions.
7. USGS-defined technological mining restrictions.
Using these data, estimates of coal resource tonnages for the Matewan
quadrangle were developed. First, however, mine planning was done manually on a printed
plot1 in equals 2,000 ft scale-of each coal seam; included on each plot was the
appropriate data (coal and parting information, mining restrictions, etc.) for that coal
seam. Then, using the GSMAP program, the mining plans for the coal seams were digitized to
produce computer data files. These data fdes, along with those supplied by the USGS, were
then used as input to the GARNET program, which does all the work of calculating the
resource tonnages.
Before resource tonnage calculations are performed, however, the GARNET program is used to
create fdes that designate different mining thickness categories and files that designate
different geographic zones within the quadrangle. The steps required to reach the resource
figures are not all the same for each mining method.
For the contour strip method-
1. The mining plan file for each coal seam consists of digitized data lines defming the toe of the fmal highwah within the seam outcrop. When this mining plan file is then combined4 with the coal seam outcrop file (fig. 6), a third file is produced that delineates the possible contour strip minable areas for the seam within the quadrangle (fig. 7).
2. Areas of restricted mining are removed from the areas delineated possible for contour strip mining (fig. 8).
The restriction files contain environmental and land use restrictions in addition to table 1 restrictions-federally funded highways (protected by a 100 ft buffer), historic sites, non-Federal public parks, and many Federal lands (e.g., national parks)-and mined-out surface and underground resources.
3. The contour strip minable seam file, when combined with the thickness category files derived from the isopach of the seam thickness, produces files delineating resources that are 12 in to 36 in thick and resources that are greater than 36 in thick (figs. 9 and 10).
4. The study area is partitioned into transportation zones in order to derive haulage costs. The center of each zone designates the source point for coal to be hauled to e)dsting preparation plants or railheads. The resource is subdivided among the transportation zones by combining the thickness files with files that delineate each of the zones. The resulting data fdes then represent the seam area that contains the contour strip minable resources, divided by thickness categories and subdivided by transportation zones (fig. 11). The figure shows aff zones and their resources as one map. In practice, however, each zone and its associated resource is a single data file, or map.
4Combine is the GARNET operation in which areas are added or subtracted.
For the continuous miner method-
1. The mining plan file for each coal seam consists of digitized data lines defining the possible underground development and production panels within the quadrangle (fig. 12)
2. Previously determined mining thickness categories of 24- to 42-in, 42- to 72-in, 72- to 96-in, and greater than 96-in seam thicknesses are used to divide the continuous miner resources.
3. The restrictions to underground mining, and thus the resources that must be eliminated, include a square (200 ft on each edge) around oil and gas wefls and coal that is within 40 ft of a sub . acent or superjacent seam of greater thickness. Mined-out sections of the quadrangle are not part of the data files-they are incorporated as a restriction during the mine planning and digitizing.
4. As with the surface mining resources, the continuous miner resources, already divided. by thickness categories, are subdivided by transportation zones (fig. 13). The resulting data files represent the part of the seam cont ' 'ng the minable continuous miner resources.
For the longwall mininig method-
1. The mining plan file for each coal seam consists of digitized data lines that outline areas in which the panels and supportive mining would be situated within the quadrangle (fig. 14).
2. Longwall resources are divided into 42- to 72-in, 72to 96-in, and greater than 96-in thicknesses.
3. Currently, afl enviromental, land use, and technological restrictions are accounted for during the mine planning; therefore, these restrictions do not need to be subsequently subtracted.
4. The resources, previously divided by thickness categories, are subdivided by transportation zones. The resulting data files then represent the portion of the seam containing the minable longwar resources (fig. 15).
Figure 16 shows all three methods of mining in a seam,
having the coal thickness isopach lines superimposed. When the resource data files for
each of the three mining methods have been created, they are checked for integrity using
GARNET. The resulting minable resources are compared with the USGS Coal Availability
results, to check for relative compatibility, and the data converted from the UNIX
computer operating system into the MS-DOS operating system. If compatibility is
established, the data is imported directly into a Lotus 1-2-3 5
version 2.2, computer program called COALVAL.
5Reference to specific trade names or manufacturers does not imply
endorsement by the U.S. Bureau of Mines.
The recoveries and costs derived during the study were calculated using
COALVAL, a menu-driven interactive spreadsheet program that was written by Bureau
personnel for use in this study. The program produces a cost analysis of coal reserves
that have previously been mine planned. COALVAL uses cost models that were developed
in-house for use in the Appalachian coal fields; the models are being modified for use in
other areas of the country as the Bureau's Coal Recoverability program matures.
COALVAL allows mine evaluators to input their reserve data and the various production,
operating, and cost variables that pertain to their property. The software can evaluate up
to 25 seams, each to be mined with up to 5 different mining methods, within a given area.
Summary spreadsheets are produced for each mining method-seam combination. The end result
of the evaluation is the cost per ton to mine the reserve.
The advantage of using COALVAL is that, relative to hand generating numbers, it quickly
and accurately costs mine-planned coal reserves. Prior to the development of the software
package, 15 person-weeks were required to cost the Matewan quadrangle reserves using the
cost models that were eventually incorporated into COALVAL. The same result can now be
accomplished in a matter of a day or two with COALVAL.
The computer operations for the Coal Recoverability study require workstations with-
1. A large (19-in), high-resolution monitor: GARNFT is a graphics program that displays and allows editing of maps.
2. Large data storage capacity: The division and subdivision of resource fdes creates data representing afl mining types, thickness categories, and transportation zones@ver 800 fdes for the 21 Matewan quadrangle seams. The fdes may range in size from a few hundred bytes to several hundred thousand bytes and require 10 to 15 megabytes of disk space per seam. The Matewan quadrangle files required 315 megabytes of disk space and additional megabytes of memory were needed for the resident software programs and the UNIX working system.
3. Networking: The computei- network link (Geonet) allows access to the NCRDS data and USGS software in Reston. Local workstations are networked together to share software, disk storage, and peripheral equipment.
4. Back-up facilities: Workstations have one-quarter inch tape drives for backing up current computer files and for archiving the completed work.
5 ' Digitizing facilities: Digitizing is done on a PC-type computer using GSMAP. The data fdes are transferred to the workstations via the network.
6. Output devices: A dot matrix printer is used for printing the output fdes from GARNET; a 42-in plotter, which allows reproduction of large maps to scale, is used to make work copies of the seam data and mine plans. A laser printer, of graphics quality, is connected to the network to provide the screen dump capability of making work and archive copies on letter size pages.
Resource recovery models were developed on COALVAL spreadsheets for each
mining method used in the quadrangle. The Matewan quadrangle includes contour strip
mining, underground room-and-pillar mining, longwaff mining, and auger mining. GARNET
output data files are reformatted and transfeffed to the spreadsheets.
Minable resource information imported from the GARNET calculations are (1) the haulage
zone, (2) acres of minable resource by zone and thickness, (3) tons of coal by zone and
thickness, and (4) tons of parting by zone and thickness. The spreadsheets calculate the
recoverable resources and operating costs associated with those resources (tables 8-11). The tons of out-of-seam dilution were separated by
mining height and mining method. Tonnage calculations use the same density for dilution
asfor in-seam parting (assumed to be 2,400 tons/acre-foot)6 The
recovery model adds the coal resource tons, in-seam parting tons, and out-of-seam dilution
tons (mining dilution), applies the appropriate mining recovery factor, and calculates a
ROM coal recovery for each mining method (tables 8-11).
Coal was considered minable until the in-seam parting thickness equalled 50% of the total
seam thickness. Although this assumption is not based on a physical m' ' g limitation, it
is reflected in the costing models because such an in-seam parting thickness is a cost
limitation. Several mine evaluations, in and around the quadrangle, indicated that the
ratio of ROM coal tonnage to waste tonnage was only slightly higher than 1:1.
The next recovery calculation deals with cleaning (washing) the ROM product to produce a
salable product. The preparation plant recovery, and the resulting salable product, are
calculated for each haulage zone by mining method and seam thickness category (tables 8-11). Review of the central Appalachian coal quality data,
modern coal preparation plants, and coal spot and contract markets provided a basis for
the models. The following items were of considerable importance:
1. Coal washing did not significantly improve the sulfur content in the salable product (i.e., most of the sulfur was tied up in the organic material, not as pyrite; therefore, it could not be easily removed).
2. Most of the modern preparation facilities examined had processes that would recover an average of 94% of the available coal from the ROM tonnage and would substitute only 6% of the parting material for coal in the fmal washed product. These average figures were incorporated into the recovery model because definitive density and washability information was not available in sufficient detail for most seams in the quadrangle.
3. The assumed cut-off between washed and directshipped coal was based on an ash content of 9%, or greater, in the ROM product. In-seam parting material and out-of-seam dilution were assumed to be 100% ash. Parting material density was assumed to be 2,400 tons/ acre-foot. Coal was assumed to have a density of 1,800 tons/acre-foot.
The total recoverable resource is calculated by subtracting from the total
original resource (including coal, inseam parting, and mining dilution) previously
mined-out and/or abandoned resources and resources unavaflable due to environmental,
technical, and barrier restrictions.
From the minable resources for each mining method, the associated mining and washing losses are subtracted. This results in the amount of recoverable resource that is available for sale, with no quality or mining cost restrictions included. Table 12 exhibits the results of the recovery calculations for one seam. The recoverable resourcesby haulage zone, mining method, and seam thickness categories are then used in the mine costing models to generate the operating costs for these recoverable coal tonnages.
Cost evaluation of the Matewan quadrangle coal resource was developed for a prefeasibility-type estimate. Economic evaluations (in-house) of mines in and around the study area were used for reference to mining methods, recovery rates, mining equipment, labor, operating costs (through the train-truck loadout), capital costs, and taxes. Industry, government, and consulting companies were used as sources of information to supplement cost figures. Costing models for underground, contour-strip, and auger mining operations were developed for different seam thicknesses; these various costing models were then integrated with local transportation (truck), depreciation, and recovery models into one interactive model-COALVALRecoverable resources are bracketed by COALVAL into different operating cost ranges: less than $25/ton, $25 to $30/ton, $30 to $40/ton, $40 to $50/ton, and over $50/ton. The costing format and summaries are developed on spreadsheets and can be updated by changing the default values associated with selected items.
Once the coal resource was categorized by seam thickness, haulage zone, and mining type, the costing models were applied to determine the economically recoverable resource. The mining type (contour strip, auger, continuous miner, or longwall) dictates the type of equipment to be used, whereas the seam thickness dictates the mining type and, therefore, the cost. The authors found that for the surface operations two loader spreads use the equipment, regardless of seam thickness; however, tile productivity will be different. The equipment list and depreciation schedule for a two-spread contour strip mine is shown in table 13. As mentioned previously, underground production is affected by mining height; consequently, continuous miner operations were divided into four different height ranges, and longwau operations were divided into three different height ranges. The type and size of equipment for selected mining heights is listed in tables 14, 15, and 16.
Once the mining type and the equipment to be used were finamd, the costs and depreciations associated with the equipment were developed. This procedure was done for each mining thickness category using that mining method to develop a total cost (fo.b., rail car) for the coal for each haulage zone. In addition, these models were designed so that they can be updated by entering new information in the assumption list at the front of each model. Only the underground costing model is presented in table 17, but a costing model was developed for each mining thickness category using each mining method.
The number of calculations involved in doing 21 seams, with as many as 16 haulage zones per seam and 9 types of mining per seam, required that an interactive model be developed. As stated previously, COALVAL accomplished this with Lotus 1-2-3 macros and linkage of spreadsheets. Furthermore, it soon became apparent as this interactive model was being developed that the model would be useful and timesaving for the Bureau, people in the coal industry, and others looking into large resource modeling projects. Therefore, a fully developed and documented model that can be used in generic situations has been published (21).
As mentioned before, the productivity of a mine is often directly related to the thickness of the coal seam being mined. This is true for both surface and underground mining operations. After deciding on the ranges of seam thickness that represented logical breaks in the productivity, the authors collected the mine equipment cost, production data for the equipment, and manning information that was based on the seam thickness involved with the model. As much as possible, the model was fashioned after a corporate cost sheet. The major factors were divided into direct and indirect operating costs, overhead, and preparation. The layout of the costing model includes a front page of 32 assumptions used in the cost calculations and a second-page summary of the cost involved in operating the particular type of mine. Following that is the heart of the cost model-the specific information and detailed calculations that are used to develop the summary on page 2 of the cost model. Since the same thinking and assumptions are used in all models, only the values change from model to model. For each mining method, the model takes into account the following:
Other items on the page 2 summary for each model are developed from the original page 1 list of assumptions, and include the following:
- Explosives: Drilling and blasting costs applicable to underground mining are not considered due to the exclusive use of continuous miners and longwall mining methods in the study. Surface drilling and blasting costs are derived from industry and Bureau publications.
- Utilities: Utility costs, such as electricity and water, are public information available from the applicable rate control agency.
- Auto expense: This cost is incurred for the use of vehicles by the supervisory personnel. Industry contacts and Bureau mine evaluations are the sources for this information.
- Insurance: Property insurance is based on a percentage per $100 property valuation. The sources for this expense are industry contacts and Bureau mine evaluations.
- Professional services: The costs of professional services address such subjects as engineering and environmental services, coal crushing and processing, accounting, project management, safety, and training. These costs were obtained from industry contacts.
- Trucking: One of the highest cost items in the mine operating cost is the transportation cost from the mine to the wash plant-loadout facilities served by the railroad. Based on present preparation plant locations within and around the Matewan quadrangle, truck haulage routes and zones were developed so that loaded trucks haul downhill. Although paved roads pass through some gaps, all travel is confined to the same watershed from which the coal is mined. Mileage calculations are made from the estimated geographic center of each zone and the cost per ton-mile applied to haul distances. Haulage costs vary from $0.65 per ton to more than $3 per ton.
- Other: Several generalized costs, including road maintenance, roof control material, coal transportation, supplies, licenses, and permits, are accounted for within various other cost categories. These costs were obtained from industry contacts and Bureau mine evaluations.
Many elements influence the cost of mining a ton of coal. Major factors to be considered are (1) geology of the resource to be mined; (2) Federal, State, and local taxes; (3) royalties; (4) payroll; (5) productivity; (6) transportation; and (7) preparation costs. The geology of the resource, as well as that of the area, not only includes the roof and floor of the seam (which directly effect the mine recovery possible from a resource and may add external dilution to the seam) but also includes the amount of parting (internal dilution found in the coal seam); this dilution directly affects the wash plant recovery that, in turn, affects the cost per ton of clean or salable coal. The mining, transportation, and cleaning of large amounts of noncoal material greatly increases the cost of mining. Federal, State, and local taxes often have little correlation to efficiency or productivity and can be a large cost consideration. Royalties can be considered a form of tax whether levied by private, State, or Federal landowners. They can often become a very large part of the total cost, as they frequently do not vary with the price of coal and usually have minimum payments required; therefore, as the coal price (or realization) or productivity decrease, the royalties become a larger percentage of the overall costs. Payroll, and the related burden rates, is usually the single largest contributor to coal mine costs. Productivity (tons of coal per employee shift) is sometimes the deciding factor between profit and loss. The term "productivity," as used in this study, includes the production of rock and shale that are mined with the coal. Transportation of the coal from the mine to the preparation facility and/or the train or barge loadout facility may account for 10% to 20% of the total operating cost. Preparation costs can have a significant impact on the cost per ton of clean coal produced, and although the Bureau has used a cost per raw ton figure in the cost models, the models allow analysis of the ROM product recovery from the plant and establish the preparation costs per clean ton of salable coal (e.g., $2 per ton for a ROM coal preparation cost, coupled with a 50% recovery in the preparation plant, equals $4 per ton for a clean coal preparation cost)-
The quality of the coal being mined does not necessarily have a direct effect on the cost to mine or wash that coal; however, the quality does affect the price that can be received for the salable product. The USGS has defmed three broad ranges of coal quality based on s@: on an as-received basis, (1) low-sulfur coal contains less than 1% sulfur; (2) medium-sulfur coal contains between 1% and 3% sulfur; and (3) high-sulfur coal contains greater than 3% sulfur (15). Compliance coal meets the current EPA standard for potential emissions from utilities burning the coal, when no sulfur dioxde (SO) reduction processes are used (23); the current standard is dependent upon the minimum allowable total emissions within an affected area and is no more than 1.2 lb of SO, per MBtu of heat input.
After the amount of recoverable coal and the quality of the coal is determined, the costing models calculate the cost to mine the coal by economic increments and quality (table 18). Figure 17 diagrammatically illustrates the procedure developed and the results produced during this study. The percentages used in the diagram all refer to the original Matewan quadrangle coal resource.
61t is recognized that roof, floor, and ir-seam parting each have different densities but due to the complexity created by trying to accommodate each density, a single standard of 2,400 tons/acre-foot was adopted.
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