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SECTION X
BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE
The effluent limitations which must be achieved by 1 July 1984,
are based on the best control and treatment technology employed
by a specific point source within the industrial category or
subcategory, or by another industry where it is readily
transferable. Emphasis is placed on additional treatment
techniques applied at the end of the treatment systems currently
employed for BPT, as well as improvements in reagent control,
process control, and treatment technology optimization.
Input to BAT selection includes all materials discussed and
referenced in this document. As discussed in Section VI, nine
sampling and analysis programs were conducted to evaluate the
presence/absence of the pollutants (toxic, conventional, and
nonconventional). A series of pilot-scale treatability studies
was performed at several locations within the industry to
evaluate BAT alternatives. Where industry data were available
for BAT level treatment alternatives, they were also evaluated.
Consideration was also given to:
1. Age and size of facilities and wastewater treatment
equipment involved
2. Process(es) employed and the nature of the ores
3. Engineering aspects of the application of various types
of control and treatment techniques
4. In-process control and process changes
5. Cost of achieving the effluent reduction by application
of the alternative control or treatment technologies
6. Non-water quality environmental impacts (including
energy requirements)
This level of technology also considers those plant processes and
control and treatment technologies which at pilot-plant and other
levels have demonstrated both technological performance and
economic viability at a level sufficient to justify investiga-
tion.
The Clean Water Act requires consideration of costs in BAT
selection, but does not require a balancing of costs against
effluent reduction benefits (see Weyerhaeuser v. Costle, 11 ERC
2149 (DC Cir. 1978)). In developing the proposed BAT, however,
EPA has given substantial weight to the reasonableness of costs
and reduction of discharged pollutants. The Agency has
considered the volume and nature of discharges before and after
497
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application of BAT alternatives, the general environmental
effects of the pollutants, and the costs and economic impacts of
the required pollution control levels. The regulations proposed
are, in fact, based on the application of what the Agency deems
to be Best Available Control Technology Economically Achievable,
with primary emphasis on significant effluent reduction
capability.
The options considered are limited only by their ability to meet
BPT Effluent Guidelines (as a minimum), technical feasibility in
the particular subcategory, and obviously extreme (high) cost.
The options presented represent a range of costs so as to assure
that affordable alternatives remain after the economic analysis.
SUMMARY OF BEST AVAILABLE TECHNOLOGY
Zero discharge limitations are established for the following
subcategories and subparts:
Iron Ore
Mills in the Mesabi Range
Mercury Ore
Mills
Cu, Pb, Zn, Au, Ag, Pt, and Mo Ores
Mills using the cyanidation process or the amalgamation
process to recovery gold or silver
Mills and mine areas that use leaching processes to
recover copper
Subcategories and subparts permitted to discharge subject to
limitations are:
Nonconventional
Pollutants Toxics
Subcategory and Subpart Controlled Controlled
Iron Ore
Mine Drainage Fe (dissolved)
Mills (physical methods) Fe (dissolved)
498
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Subcategory and Subpart
Nonconventional
Pollutants
Controlled
Toxics
Controlled
Subcategory and Subpart
Controlled
Controlled
Aluminum Ore
Mine Drainage
Uranium, Radium, and
Vanadium Ores
Mine Drainage
Mercury Ore
Mine Drainage
Titanium Ore
Mine Drainage
Mills
Dredges
Tungsten Ore
Mine Drainage
Mills
Cu, Pb, Zn, Au, Ag,
Pt, and Mo Ores
Mine Drainage (not
placer mining)
Mills (froth
flotation)
Fe, Al
COD, Ra226 (dis-
solved) Ra226
(total), U
Zn
Fe
Fe
Hg
Zn
Cd, Cu, Zn
Cd, Cu, Zn
Cd, Cu, Zn,
Pb, Hg,
Cd, Cu, Zn,
Pb, Hg,
499
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The specific effluent limitations guidelines for the subcate-
gories and subparts permitted to discharge are:
Toxic Pollutants
Copper
Zinc
Lead
Mercury
Cadmium
Daily
Maximum
mg/1
0.30
1.0 (1
0.6
0.002
0. 10
30-Day
Average
mg/1
0.15
5)* 0.5 (0.75)*
0.3
0.001
0.05
Nonconventional Pollutants
Iron (dissolved)
Iron (total)
Aluminum
COD
Radium 226 (dissolved)
Radium 226 (total)
Uranium
Daily
Maximum
mg/1
30-Day
Average
mg/1
2.0
2.0
10 (pCi/1)
30 (pCi/1)
4
1 .0
1 .0
1 .0
500
3 (pCi/1)
10 (pCi/1)
2
*Limitations applicable to mine drainage for the copper,
lead, zinc, gold, silver, platinum, and molybdenum ores
subcategory.
GENERAL PROVISIONS
Several items of discussion apply to options in more than one
subcategory. To avoid repetition, these items are discussed here
and referred to in the discussion of the options.
Relief From No Discharge Requirement
Facilities which are not allowed to discharge under
conditions may do so as a result of:
"normal"
1. An overflow or increase in volume from a precipitation
event if the facility is designed, constructed and
maintained to contain a 10-year, 24-hour rainfall design
(storm provision); or
2. If they are located in a "net precipitation" area.
These exemptions are discussed in greater detail below.
500
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Storm Provision
The Agency recognizes that relief is necessary as a practical
matter for many discharges within the ore mining and dressing
point source category during and immediately after some
precipitation events. It would be unreasonable to require
facilities to construct retention structures and treatment
facilities to handle runoff resulting from extreme rainfall con-
ditions which could statistically occur only rarely. Further, it
must be emphasized that the regulations for the ore mining and
dressing point source category do not require any specific
treatment technique, construction activity, or other process for
the reduction of pollution. The effluent limitations guidelines
limit the concentration of pollutants which may be discharged,
while allowing for an excursion from the normal requirements when
precipitation causes an overflow or increase in the volume of a
discharge from a facility properly designed, constructed, and
maintained to contain or treat a 10-year, 24-hour rainfall.
This relief applies to the excess volume caused by precipitation
or snow melt, and the resulting increase in flow or shock flow to
the settling facility or treatment facility. While there has
been criticism of the relief adopted by the Agency, the few
alternatives suggested by environmental groups and industry are
substantially less satisfactory in light of the data available to
the Agency. This is discussed in detail in the preamble to the
BPT regulation (43 FR 29771) and in the clarification of
regulations (44 FR 7953).
The general relief in the BPT regulation states that:
"Any excess water, resulting from rainfall or snow melt,
discharged from facilities designed, constructed and
maintained to contain or treat the volume of water which
would result from a 10-year, 24-hour pecipitation event
shall not be subject to the limitations set forth in 40 CFR
440." 43 FR at 29777-78, 440.81(c)(1978).
The term "ten-year, 24-hour precipitation event" is defined,
in turn, as:
"the maximum 24-hour precipitation event with a probable re-
occurrence interval of once in 10 years as defined by the
National Weather Service and Technical Paper No. 40,
'Rainfall Frequency Atlas of the U.S.,1 May 1961, and
subsequent amendments, or equivalent regional or rainfall
probability information developed therefrom." 43 FR at
29778, 440.82(d).
Under BAT, similar relief is granted: for existing sources that
are designed, constructed, and maintained to contain or treat the
maximum volume of process wastewater discharged in a 24-hour
period, including the volume which would result from a 10-year,
501
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24-hour precipitation event, or snow melt of equal volume, any
excess wastewater discharged shall not be subject to the
limitations set forth in 40 CFR 440.
In determining the maximum volume of wastewater which would
result from a 10-year, 24-hour precipitation event at any
facility, the volume must include the volume that would result
from runoff from all areas contributing runoff to the individual
treatment facility, 'i.e., all runoff that is not diverted from
the active mining area, runoff which is not diverted from the
mill area, and other runoff that is allowed to commingle with the
influent to the treatment system.
In general, the following will apply in granting relief:
1. The exemption as stated in the rule is available only if
it is included in the operator's permit. Many existing
permits have exemptions or relief clauses stating
requirements other than those set forth in the rule. Such
relief clauses remain binding unless and until an operator
requests a modification of his permit to include the
exemption as stated in the rule.
2. The storm provision is an affirmative defense to an
enforcement action. Therefore, there is no need for the
permitting authority to evaluate each tailings pond or
treatment facility now under permit.
3. Relief can be granted to deep mine, surface mine, and
ore mill discharges.
4. Relief is granted as an exemption to the requirements
for normal operating conditions (i.e., without overflow,
increase in volume of discharge, or discharge from a by-pass
system caused by precipitation) with respect to an increase
in flow caused by surface runoff only.
5. Relief can be granted for discharges during and immedi-
ately after any precipitation or snow melt. The intensity
of the event is not specified.
6. The relief does not grant, nor is it intended to imply
the option of ceasing or reducing efforts to contain or
treat the runoff resulting from a precipitation event or
snow melt. For example, an operator does not have the
option of turning off the lime feed to a facility at the
start of or during a precipitation event, regardless of the
design and construction of the wastewater facility. The
operator must continue to operate his facility to the best
of his ability.
7. Under the regulation, relief is granted from all efflu-
ent limitations contained in BAT.
502
-------
8. In general, relief can not be granted to treatment
facilities which employ clarifiers, thickeners, or other
mechanically aided settling devices. The use of
mechanically aided settling usually is restricted to
discharges which are not affected by runoff.
9. In general, the relief was intended for discharges from
tailings ponds, settling ponds, holding basins, lagoons,
etc. that are associated with and part of treatment
facilities. The relief will most often be based on the
construction and maintenance of these settling facilities to
"contain" a volume of water.
10. The term "treat" for facilities allowed to discharge
means the wastewater facility was designed, constructed, and
maintained to meet the daily maximum effluent limitations
for the maximum flow that would result from a 10-year, 24-
hour rainfall. The operator has the option to "treat" the
volume of water that would result from a 10-year, 24-hour
rainfall in order to qualify for the rainfall exemption, or
as mentioned in paragraph 9 above, the second option is to
"contain" the volume of wastewater.
11. The term "maintain" is intended to be synonymous with
"operate." The facility must be operated at the time of the
precipitation event to contain or treat the specified volume
of wastewater. Specifically, in making a determination of
the ability of a facility to contain a volume of wastewater,
sediment and sludge must not be permitted to accumulate to
such an extent that the facility cannot in fact hold the
volume of wastewater resulting from a 10-year, 24-hour
rainfall. That is, sediment and sludge must be removed as
required to maintain the specific volume of wastewater
required by the rule, or the embankment must be built up or
graded to maintain a specific volume of wastewater required
by the rule.
12. "Contain" and "maintain" for facilities which are
allowed to discharge do not mean providing for draw down of
the pool level of the facility allowed to discharge. As an
example, the volume can be determined from the top of the
stage of the highest dewatering device to the bottom of the
pond at the time of the precipitation event. There is no
requirement that relief be based on the facility which is
allowed to discharge being emptied of wastewater prior to
the rainfall or snow melt upon which the relief is granted.
The term "contain" for facilities which are allowed to
discharge means the wastewater facility's tailings pond or
settling pond was designed to include the volume of water
that would result from a 10-year, 24-hour rainfall.
13. The term "contain" for facilities which are not allowed
to discharge means the wastewater facility was designed,
503
-------
constructed, and maintained to hold, without a point source
discharge, the volume of water that would result from a 10-
year, 24-hour rainfall, in addition to the normal amount of
water which would be in the wastewater facility.
Net Precipitation Areas
The general relief or exemption for the requirement of "no
discharge of process wastewater" as promulgated for ore mining
and dressing in BPT Guidelines (43 FR 29771) states that:
"In the event that the annual precipitation falling on the
treatment facility and the drainage area contributing
surface runoff to the treatment facility exceeds the annual
evaporation, a volume of water equivalent to the difference
between annual precipitation falling on the treatment
facility and the drainage area contributing surface runoff
to the treatment facility and annual evaporation may be
discharged subject to the limitations set forth in paragraph
(a) of the section." Paragraph (a) refers to limitations
established for mine drainage.
Relief for net precipitation areas is included in the BAT
regulation.
Commingling Provision
The general provision as promulgated for ore mining and dressing
in the BPT Guidelines (43 FR 29771) states that:
"In the event that waste streams from various subparts or
segments of subparts in part 440 are combined for treatment
and discharge, the quantity or quality of each pollutant or
pollutant property in the combined discharge that is subject
to effluent limitations shall not exceed the quantity or
quality of each pollutant or pollutant property that would
have been discharged had each waste stream been treated
separately. The discharge flow from a combined discharge
shall not exceed the volume that would have been discharged
had each waste stream been treated separately."
This consideration is also appropriate for BAT and the regulation
states:
For existing sources which as of the date of this proposal
have combined for treatment waste streams from various
subparts or segments of subparts in Part 440, the quantity
and quality of each pollutant or pollutant property in the
combined discharge that is subject to effluent limitations
shall not exceed the quantity and quality of each pollutant
or pollutant property that would have been discharged had
each waste stream been treated separately. The discharge
flow from a combined discharge shall not exceed the volume
504
-------
that would have been discharged had each waste stream been
treated separately.
BAT OPTIONS CONSIDERED FOR TOXICS REDUCTION
As discussed in Section VII, many toxic pollutants found in this
category are related to TSS (that is, as TSS concentrations are
reduced during treatment, observed concentrations of certain
toxic metals are also reduced). In order to remove these toxics,
suspended solid removal technologies can be used. The technolo-
gies are secondary settling, coagulation and flocculation, and
granular media filtration. They are applicable throughout the
category for suspended solids reduction and associated toxic
metals reduction, and are discussed ' here to avoid repetition
during description of options. Dissolved metals are not
controlled further by physical treatment methods or additional
suspended solids removal.
Secondary Settling
This option involves the addition of a second settling pond in
series with the existing pond (as described in Section VIII).
The technique is used in many ore subcategories. The most
prevalent configuration is a second pond located in series with a
tailings pond.
Examples of the use of secondary and tertiary settling ponds can
be seen at lead/zinc Mills 3101, 3102, 3103 and at Mill 4102
(Pb/Zn/Au/Ag). This last facility uses a secondary pond to
achieve an effluent level of 4 mg/1 TSS, as determined during
sampling (Reference 1). Secondary settling ponds (sometimes
called polishing ponds) are also used in settling solids produced
in the coprecipitation of radium with barium salts at uranium
mines and mills. (See Section VIII, End-of-Pipe Techniques,
Secondary Settling; and Historical Data Summary, lead/zinc Mills
3101, 3102, 3103, and 4102.)
Coagulation and Flocculation
Chemically aided coagulation followed by flocculation and
settling is described in Section VIII. It is used by facilities
in several subcategories of the industry for solids and metals
reduction.
At Mine/Mill 1108, the tailing pond effluent is treated with alum
followed by polymer addition and secondary settling to reduce TSS
from 200 mg/1 to an average of 6 mg/1. At Mine 3121, polymer
addition has greatly improved the treatment system capabilities.
A TSS mean concentration of 39 mg/1 (range 15 to 80) has been
reduced to a mean of 14 mg/1 (range 4 to 34), a reduction of 64
percent. Similarly, polymer use at Mine 3130 reduced treated
effluent total suspended solids concentrations from a mean of 19
mg/1 (range 4 to 67 mg/1) to a mean of 2 mg/1 (range of 1 to 6.2
505
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mg/1). It should be pointed out that these effluent levels are
attained by the combination of settling aids and a secondary
settling pond (Section VIII).
Granular Media Filtration
This option uses granular media such as sand and anthracite to
filter out suspended solids, including the associated metals (as
discussed in Section VIII). This technology is used at one
facility (Mine 6102) and has been pilot tested at other mines and
mills and is used in other industry categories.
Granular-media filtration can consistently remove 75 to 93
percent of the suspended solids from lime-treated secondary
sanitary effluents containing from 2 to 139 mg/1 of suspended
solids (Reference 2). In 1978, lead/zinc
Mine/Mill/Smelter/Refinery 3107 was operating a pilot-scale
filtration unit to evaluate its effectiveness in removing
suspended solids and nonsettleable colloidal metal-hydroxide
floes from its wastewater treatment plant. Granulated slag was
used as the medium in some of the tests. Preliminary data
indicate that the single medium pressure filter was capable of
removing 50 to 95 percent of the suspended solids and 14 to 82
percent of the metals (copper, lead and zinc) contained in the
waste stream. Final suspended solids concentrations which have
been obtained are within the range of 1 to 15 mg/1. Optimum
filter performance was attained at lower hydraulic loadings (2.7
A full-scale multi-media filtration unit is currently in
operation at molybdenum Mine/Mill 6102. The filtration system is
used as treatment following settling (tailing pond), ion exchange
(for molybdenum removal), lime precipitation, electrocoagulation,
and alkaline chlorination. Since startup in 1978, the filtration
unit has been operating at a flow of 63 liters/second (1000 gpm)
and monitoring data show TSS reductions to an average of less
than 5 mg/1. Zinc removal and iron reduction have also been
achieved (see Treatment Technology - Section VIII).
A pilot-scale study of mine drainage treatment in Canada has also
demonstrated the effectiveness of filtration (Section VIII).
Polishing of clarifier overflow by sand filtration resulted in
reduction of the concentration of lead and zinc (approximately 50
percent) and removal of iron (approximately 40 percent) and
copper.
In addition to the above, a full-scale application of slow sand
filters is employed at iron ore Mine/Mill 1131 to further polish
tailing pond effluent prior to final discharge.
Besides the application at the various facilities described
above, a series of pilot-scale tests was performed at a number of
facilities in the ore mining category as part of the investi-
gation of BAT technologies described. These studies were con-
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ducted at Mine/Mill 3121, Mine/Mill/Smelter/Refinery 3107, Mill
2122 (two studies on tailing pond effluent), Smelter/Refinery
2122 (wastewater treatment plant), Mine/Mi 1I/Smelter/Refinery
2121, Mine 3113, Mine 5102, Mill 9401, Mill 9402 (two studies)
(Reference Section VIII). In each case, filtration (among other
technologies) was evaluated and produced average effluent levels
of TSS consistently below 10 mg/1, and usually below 5 mg/1 on an
average basis.
Partial Recycle
This option consists of the recycle and reuse of mill process
water (not once-through mine water used as mill process water).
One of the principal advantages of recycle of process water is
the volume of wastewater to be treated and discharged is reduced.
Although initial capital costs of installation of pumps, piping,
and other equipment may be high, these are often offset by a
reduction in costs associated with the treatment and discharge.
Many facilities within this industry practice partial recycle
including lead/zinc Mills 3105 (67 percent), 3103 (40 percent),
3101 (all needs met by recycle), gold Mill 4105 (recycle of
treated water), molybdenum Mill 6102 (meets needs of mill),
nickel Mill and Smelter 6106, vanadium Mill 6107, and titanium
Mill 9905. In-process recycle of concentrate thickener overflow
and/or filtrate produced by concentrate filtering is practiced by
a number of flotation mills including 2121, 3101, 3102, 3108,
3115, 3116, 3119, 3123, and 3140. In addition, Mills 2120, 1132,
6101, and 6157 employ thickeners to reclaim water from tailings
or settling ponds prior to the final discharge of these tailings
to tailings ponds.
The practices described above are beneficial with respect to
water conservation and recovery of metals which might be lost in
the wastewater discharge. These practices are also significant
with respect to wastewater treatment considerations. The in-pro-
cess recycle of concentrate-thickener overflow and/or filtrate
produced by concentrate filtering reduces the volume of waste-
water discharged by 5 to 17 percent at mills which employ these
practices. Likewise, those mills which reuse water reclaimed
from tailings reduce both new water requirements and the volume
discharged by 10 to 50 percent. The advantage of any practice
which reduces the volume of wastewater discharged can be viewed
in terms of economy of treatment and enhancement of treatment
system capabilities (i.e., increased retention time of existing
sedimentation basins).
The use of mine drainage as makeup in the mill is a practice that
also deserves mention here as a method of reducing discharge
volume to the environment. A large number of facilities in the
ore mining and dressing point source category employ this
practice (see Section VIII).
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In general, there are four benefits resulting from adoption of
this practice. They are:
1. Recovery of raw materials in processing;
2. Conservation of water;
3. Reduction of discharge to tailings ponds; and
4. Increase in performance of tailings ponds.
Implementing recycle within a facility or treatment process may
require modification. Modification will be specific to each
facility and each operator will have to make his own determina-
tions.
100 Percent Recycle - Zero Discharge
This option consists of complete recycle and reuse of process
water with no resulting discharge of wastewater to the environ-
ment. Many facilities in the industry have demonstrated that
total recycle of process water is technically feasible. All of
the iron ore mills in the Mesabi Range have demonstrated the
viability of this option. Total recycle systems are also
demonstrated by iron ore Mill 1105, rare earth Mill 9903 and
mercury Mill 9201. Forty-six mills using froth flotation in the
Copper, Lead, Zinc, Gold, Silver, Platinum and Molybdenum Ores
Subcategory presently achieve zero discharge including 31 copper,
five lead/zinc, five primary gold, one molybdenum and four
primary silver mills.
There are two methods of water reclamation that are practiced in
a number of mills. They are in-process recycle and end-of-pipe
recycle. In-process recycle may involve recycle of overflow from
concentrate thickeners, recycle of filtrate from concentration
filters, recycle of spilled reagents or any combination of these.
End-of-pipe recycle involves recycle of overflow from a tailings
thickener or recycle from the tailings pond itself.
For facilities practicing mining and milling, it can be argued
that in many cases the combined treatment of mine and mill waste-
water is beneficial from a discharge standpoint, however, the
feasibility of combining the mine and mill streams will depend on
the magnitudes of:
1. The flow of mine drainage; and
2. The process water makeup flow required for the mill.
In order to achieve zero discharge at many mills, it will be
necessary to treat mine water separately and use part of it as
makeup water in the mill.
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In comments to EPA concerning existing facilities, some
facilities claimed that zero discharge by recycle would
contribute to the buildup of reagents in the process water. It
was further alleged that this buildup would interfere with the
metallurgy of the process. To date, no data have been submitted
demonstrating this contention. The cost of preventing runoff
from entering tailings or sedimentation ponds and geographical
locations in areas which have net precipitation are also concerns
with this option for existing sources.
SELECTION AND DECISION CRITERIA
Summary of_ Pollutants to be Regulated
In Section VII, Selection of Pollutant Parameters, the effluent
data obtained during sampling and analysis for each of the 129
toxic pollutants were reviewed by subcategory and subpart for
further consideration in regulation development. In summary, all
114 of the toxic organic, and six of the toxic metal pollutants
were excluded from further consideration under provisions in
Paragraph 8 of the Settlement Agreement as shown below.
Toxic Pollutants
129
Organics - 87 (Not Detected)
Organics - 17 (Detected at levels below EPA's
nominal detection limit)
Organics - 9 (Detected at levels too low to be
effectively treated)
Organic - 1 (Uniquely related to the facility
in which it was detected)
Metals - 6 (Detected at levels too low to be
effectively treated)
9 (Remaining for consideration)
7 Toxic Metals (arsenic, copper, lead, zinc,
cadmium, mercury and nickel) Asbestos and
Cyanide
The seven toxic metals, asbestos and cyanide were considered for
regulation in subcategories and subparts where these pollutants
were detected during sampling and analysis and were not excluded
under Paragraph 8 as discussed in Section VII. Chemical oxygen
demand, total iron, dissolved iron, total radium 226, dissolved
radium 226, aluminum, and uranium, are regulated in the
subcategories and subparts in which they were regulated under
BPT.
Subcategories and Subparts in Which Toxic Pollutants Were Not
Detected Are Excluded Under Paragraph 8_ of_ the Settlement
Agreement
There were subcategories and subparts in which all of the toxic
pollutants were excluded from further consideration in regulation
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development (refer to Table VII-2, Pollutants Considered for
Regulation). These include:
1. Iron ore mine drainage and mill process wastewater (not
in the Mesabi Range);
2. Aluminum mine drainage;
3. Titanium mine drainage (lode ores); and
4. Titanium mines/mills employing dredging of sand
deposits.
Consequently, for these subcategories and subparts, BAT effluent
limitations are the same as BPT effluent limitations since there
are no toxic pollutants to be controlled.
Subcategories and Subparts Which Were Not Permitted to Discharge
Under BPT
No discharge of wastewater was specified for facilities in the
following subcategories and subparts under BPT:
1 . Iron Ore Mills in the Mesabi Range;
2. Copper, Lead, Zinc, Silver, Gold, Platinum and
Molybdenum Mines and Mills that leach to recover copper;
3. Gold Mills that use cyanidation; and
4. Mercury Mills.
Facilities in these subcategories and subparts have achieved the
goal of the Clean Water Act and no additional reduction of toxics
is possible. Therefore, the BAT effluent limitations are the
same as under BPT.
Subcategories and Subparts Where BAT Limitations Are Developed
There were subcategories and subparts in which some of the toxic
pollutants were detected and are not excluded from regulation
development. These include:
1. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum mine drainage, mill process water from facilities
using froth flotation, and placer mines;
2. Tungsten mine drainage and mill process water;
3. Mercury mine drainage;
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4. Uranium mine drainage; and
5. Titanium mill process water.
Recycle was considered as an option for froth flotation mills in
the copper, lead, zinc, gold, silver, platinum, and molybdenum
ores subcategory. This option was rejected for froth flotation
mills because of the extreme costs that would be incurred during
retrofit, the downtime required to retrofit existing equipment
and the impact of changing the metallurgical process.
Recycle was also considered as an option for placer mines
recovering gold. The placer mining industry consists primarily
of small operations located in remote areas in Alaska. Placer
mining involves recovery of gold and other heavy mineral deposits
by washing, dredging, or other hydraulic methods. Chemical
reagents are not used in the processing of the deposits. For
this reason, the pollutants of primary concern are suspended or
settleable solids which may result during recovery.
Arsenic and mercury were found in placer effluents during recent
studies because (1) arsenic occurs naturally in abundance in many
areas of Alaska; and (2) mercury has been used extensively in the
past by placer miners for recovery of gold in sluiceboxes, and
mercury residuals are undoubtedly present in old deposits
presently being reworked by modern day miners. Results of this
study indicate that effective removal of the total suspended and
settleable solids by settling also resulted in effective removal
of arsenic and mercury.
At a few placer mines it may be technically feasible to recycle
water for reuse in sluicing gold-bearing sediments. However, the
location of most of the operations, the fact that electric power
is not available to run pumps and the magnitude of the costs and
energy requirements mitigate against this practice. As a result,
EPA has selected settleable solids limitations based on settling
ponds as the means for controlling discharges from placer mining
operations. The choice of settleable solids frees the operators
from having to ship samples from remote locations to laboratories
for analysis. The analytical method is undemanding, inexpensive,
short-term duration test that can be performed by large and small
operators alike.
The settelable solids data from placer mining facilities included
two separate studies of existing placer mines in Alaska and other
studies performed by EPA and by departments of the State of
Alaska. However, the actual data for effluent from existing
settling ponds associated with gold placer mines is limited
because many of the mines, including mines in the data base, have
no settling facilities. Of the remaining mines which have
settling ponds, it was identified that the majority, if not all,
of the existing ponds for which we have data, were undersized,
filled with sediment, short circuited, or otherwise poorly
511
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operated to remove settlable solids from the wastestreams of
placer mines. The data from well constructed, operated, and
maintained settling ponds is limited to demonstration projects
and a few existing settling ponds which may not be truly
representative of gold placer mining operations (e.g., mines
located outside of the boundries of streams or floating dredge
operations).
Cost comparisons for two treatment technologies (primary settling
followed by secondary settling and primary settling with
flocculation) were perfomred including the subsequent cost per
ton of ore mined. However, no economic analysis was perfomred
for the gold placer mining subpart because no data are available
that would enable the Agency to perform cash flow analysis of
placer mine operations.
Limitation for gold placer mines are reserved in the proposed BAT
rulemaking in the absence of information regarding the economic
impact of regulating gold placer mines and to allow the Agency to
seek additional data on the effluent from well operated settling
ponds assocaited with gold placer mines.
The method used to compute the achievable levels for the toxic
metals is summarized below and presented in greater detail in
Supplement B. The data obtained during sampling and analysis as
well as that supplied by industry were reviewed and effluent
levels achievable were computed for each toxic metal considered
for regulatory development in Section VII. As discussed in
Section VII, TSS removal technologies also remove metals, so the
following TSS removal measures were considered for metal removal:
1. Secondary settling;
2. Coagulation and flocculation; and
3. Granular media filtration.
Eighteen facilities throughout the ore mining and dressing
industry were identified as using multiple settling ponds; 14
facilities using coagulation and flocculation; and one facility
using granular media filtration. The entire BAT and BPT data
base was searched and screened to obtain 17 facilities with data.
Of these 17 facilities, seven were eliminated because it was
believed that they were not operated properly (e.g., short
circuiting in the settling ponds was observed) or no raw
(untreated) wastewater data were available to compare with
treated effluent.
The facility treated effluent mean values were ranked for each of
the 10 remaining facilities for each pollutant from largest to
smallest. Since each facility used only one of the candidate BAT
treatment technologies, the facility mean also represented a
treatment technology mean value. When examining the ranked mean
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values, it was observed that mean values for facilities using
secondary settling bracketed those for facilities using
flocculation and granular media filtration. This variation
indicated that the differences between facilities were greater
than the differences between treatment technologies. Possibly,
differences existed between the true performance capabilities of
the treatment technology; however, on the basis of available
data, one cannot discern such differences.
The 10 facilities were then further reduced to six by eliminating
facilities whose raw (untreated) waste contained low pollutant
concentrations. This was done to ensure that only those
facilities which demonstrated true reduction would be included in
the analysis. Data for a particular pollutant were excluded if
the median raw wastewater concentration was less than the average
facility effluent concentration of any other facility. Of the
six facilities, five use secondary settling and one uses granular
media filtration. Since there were no discernable differences in
the levels achievable by the three technologies (based on
available data), the least costly alternative was selected for
establishing effluent limitations, secondary settling.
Achievable levels were computed by using the average of the
facility averages for each pollutant to represent, the average
discharge. The data used were from the five facilities using
secondary settling (two copper, two lead/zinc, and one silver)
that remained following the screening procedures described above.
The data base indicates that within-plant effluent concentrations
were approximately log normally distributed. The 30-day average
maximum and daily maximum effluent limits were determined on the
basis of 99th percentile estimates. The 30-day limits were
determined by using the central limit theorem. The achievable
levels computed for each of the metals and TSS are shown below:
30-Day Daily
Average Maximum
Arsenic 0.01 0.05
Cadmium 0.005 0.01
Copper 0.05 0.20
Lead 0.04 0.14
Mercury 0.001 0.002
Nickel 0.10 0.40
Zinc 0.20 0.80
TSS* 10 25
*TSS limitations were computed, but TSS
would be limited under BCT.
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The limitations derived from the data analysis for some pollutant
metals were more stringent than the BPT limitations.
Having computed achievable levels for the candidate BAT
technologies, EPA then completed an environmental assessment,
which analyzed the environmental significance of toxic pollutants
currently discharged from facilities in this industry and also
those toxic pollutants known to be discharged from this industry
at BPT and expected to be discharged based on the computed
achievable levels. The basis for determining the environmental
significance of toxic pollutants in current discharges is a
comparison of average plant effluent concentrations with Ambient
Water Quality Criteria (WQC) for the protection of human health
and aquatic life published by the EPA's Criteria and Standards
Division (CSD) in November 1980. Because WQC for the protection
of aquatic life were not developed for all of the Section
307(a)(1) toxic pollutants, the average plant effluent concentra-
tions for pollutants lacking these WQC are compared with pollu-
tant-specific toxicity data reported in the Ambient Water Quality
Criteria Documents. The environmental significance of toxic
pollutants in post-BAT discharges is determined by comparing the
achievable levels with WQC or, for those pollutants lacking WQC
for the protection of acquatic life, with EPA toxicity data.
Based on a review of the sampling and analysis data available for
this industry, the only environmentally significant pollutants
after applying the median dilution from the average receiving
stream flow available (to this industry) are cadmium and arsenic.
The concentration of cadmium currently being discharged from this
industry (BPT) is the lowest of any industry known to discharge
cadmium. In addition, the additional BAT reductions are small
relative to the levels present in raw (untreated) waste streams.
In preparing the environmental assessment, the Agency also
compared raw waste mass loadings to those of BPT and those
expected by achievable levels. It was found that the industry's
current discharge is less than 10 percent of the industry's raw
waste load. This is due to the installation and proper
maintenance of the Best Practicable Technology at many plants.
In other information and data available to the Agency at the time
this Development Document was written, the BAT limitations
proposed or being considered for metals in discharges from other
industries are less stringent than those promulgated for BPT in
the Ore Mining and Dressing Industry. In examining limitations
for Coil Coating, Metal Finishing and Inorganic Chemicals, the
following observations may be made with respect to the
limitations:
1. Based on the use of similar technology (lime precipi-
tation and settling), proposed BAT limitations on copper,
cyanide, mercury and zinc for Coil Coating are less
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stringent than those promulgated as BPT for Ore Mining and
Dressing. Cadmium and lead limitations are more stringent.
2. Based on the use of similar technology (lime precipi-
tation and settling), proposed BAT limitations for the
following pollutants being considered for Inorganic
Chemicals are less stringent than in Ore Mining BPT in the
following subcategories.
- Sodium Dichromate - zinc (copper, lead, cyanide,
and mercury are not regulated)
cadmium
Copper Sulfate - copper, cadmium and zinc (mercury and
cyanide are not regulated; lead is more stringent)
Nickel Sulfate - copper, cadmium, and zinc (mercury and
cyanide are not regulated; lead is more stringent)
Sodium Bisulfite - copper, mercury, and zinc (cadmium and
cyanide are not regulated; lead is more stringent)
Sodium Hydrosulfite - copper, lead and zinc (cadmium,
cyanide and mercury are not regulated)
- Aluminum Fluoride - copper (cadmium, cyanide, lead, and
mercury are not regulated; zinc is more stringent)
Titanium Dioxide - cadmium, copper and zinc (cyanide and
mercury are not regulated; lead is more stringent)
3. Based on the use of similar technology (lime precipita-
tion and settling), limitations for the Common Metals
Subcategory in Metal Finishing on copper and zinc are less
stringent than those promulgated for Ore Mining and
Dressing. Cadmium and lead limitations are more stringent.
EPA also considered the solubility of compounds in which the
toxic metals are found in the industry effluent. The metals
present in the wastewater from ore milling are found primarily in
the ore and gangue discharged by the mill. The metals present in
mine drainage similarly consist primarily of ore and gangue. The
metals in ore and gangue are primarily sulfides and to a lesser
extent oxides. The metals as mined generally exist as sulfides:
as example, chalcopyrite (CuFeS2), bornite (Cu5FeS4),
(CuS), and chalcocite (Cu2S) for copper; galena (PbS!
sphalerite (ZnS) for zinc; and argentite (Ag2S),
(Ag3AsS3) and stephanite Ag5S4Sb for sliver. The metals
exist as
for copper
zinc; and
example,
(Zn2Si04)
covellite
for lead;
proustite
may also
oxides: as example, tenorite (CuO) and cuprite (Cu20)
• pyromorphite Pb5Cl(P04)3 for lead; zincite (ZnO) for
wulfenite (PbMo04) for molybdenum or silicates, as
hemimorphite, Zn4 (Si207
for zinc (Reference 5).
(OH
H0
The oxides
and willemite
and silicates are
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less soluble than the sulfides and the sulfides are far less
soluble than the hydroxides as shown in the table below.
Solublility Products of Toxic Metals
Solubility Product
Toxic Metal Metal Hydroxide Metal Sulfide
Cadmium, Cd
Copper, Cu
Lead, Pb
Mercury, Hg
Zinc, Zn
13.6
18.6
16. 1
25.4
15.7
26.1
35.2
26.6
52.2
25.2
Source: Development Document for Effluent Limitations
Guidelines and Standards for the Inorganic
Chemicals Manufacturing Point Source Category,
EPA 440/1-79/007, June 1980.
By contrast, in many other industrial point source categories
including, for example, the coal coating, metal finishing,
inorganic chemical categories, toxic metals are introduced with
raw waste as dissolved species. The toxic metals in those
industries are present in the wastewater either in the dissolved
form or as metal hydroxide. As indicated by the above table
metal hydroxides are of comparatively high solubility.
The metal hydroxides (solids) present in the wastewater of these
other industries have a greater potential of redissolving into
solution and being available to the environment to be assimilated
thereby posing a greater danger to human and aquatic species.
The metals (solids) present in the wastewater from ore milling
consist primarily of the minerals associated with the ore and
gangue discharged by the mill. The metals (solids) present in
mine drainage similarly consists of ore and gangue. The metals
in the ore and gangue are generally the sulfides of the metal and
are of relatively low solubility.
After considering the environmental assessment, the BAT
limitations proposed for other industries, the physical form the
toxic metals, and economic factors, the Agency has concluded that
nationally applicable regulations based on secondary settling or
any of the other candidate BAT technologies are not warranted in
the Ore Mining and Dressing Point Source Category.
Cyanide Control and Treatment
As discussed in Section VIII, cyanide compounds are used in froth
flotation process of copper, lead, zinc, and molybdenum, ores.
In addition, the cyanidation process is used for leaching of gold
and silver ores. Consequently, residual cyanide is found in mill
tailings and wastewater streams from these mills. Cyanide is
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also found in low concentrations in mine drainage at facilities
which backfill mine stopes with the sand fraction of mill
tailings.
Of the control and treatment technologies available for cyanide,
consideration was given to the following options: in-process
control, chemical oxidation (alkaline chlorination, hydrogen
peroxide oxidation, ozonation), and natural oxidation and the
incidental removal occurring in existing treatment systems
(tailing ponds). These options were judged to be most applicable
to the high flow volume and comparatively low concentrations of
cyanide in the wastewater streams typical in this category.
Another alternative which was considered was the substitution of
other reagents for cyanide compounds in froth flotation
processes. Bench-scale testing indicated that this alternative,
although technically feasible, would require extensive testing in
actual production of circumstances with specific ores. In
addition, it would be difficult in these cases to predict
downtime, loss of recovery (if any), and costs associated with
process modifications.
Alkaline Chlorination
This method was described in
operating cost assumptions
Basically, oxidation of cyanide
accomplished by infusion of
stream at a pH greater than 10,
detail in Section VIII, while
are outlined in Section IX.
by alkaline chlorination may be
gaseous chlorine into the waste
or by the addition of sodium
hypochlorite (NaOCl) as an oxidant along with an alkali such as
sodium hydroxide (NaOH). The alkali achieves pH adjustment and
precipitation of metal hydroxides formed from the breakdown of
metal-cyanide complexes.
Pilot-scale tests of alkaline chlorination treatment at Mill 6102
showed reduction of effluent cyanide concentrations from 0.19
mg/1 to less than 0.1 mg/1 at pH values greater than 8,8. In
addition, Mill 3144 achieved reduction of effluent cyanide
concentrations to an average of 0.18 mg/1 from 4.72 mg/1
following the installation of a full-scale alkaline-chlorination
treatment system.
Ozonation
Oxidation of cyanide by ozonation is also accomplished at ele-
vated pH (9 to 12). Copper appears to act as a catalyst in this
process, which suggests that waste streams containing copper
cyanide complexes may be treated more effectively by ozonation.
Pilot-scale testing of ozonation at Mill 6102 showed reduction of
cyanide concentration from 0.55 mg/1 to less than 0.1 mg/1 at pH
greater than 7.4.
Hydrogen Peroxide Treatment
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Hydrogen peroxide (H202) has also been tested on a limited basis
as an oxidant for cyanide treatment in milling wastewater
streams. This process also requires an alkaline pH and can be
enhanced by a copper catalyst. Mill 6101 has achieved approxi-
mately 40 percent removal of cyanide during periods of elevated
effluent levels (up to approximately 0.09 mg/1) by hydrogen
peroxide oxidation.
Process Control
One characteristic of the froth flotation process which poten-
tially affects effluent wastewater quality is the latitude avail-
able to the mill operator at the upper end of the dosage applica-
tion spectrum. That is, while the addition' of less than the
necessary quantities of cyanide reagent may lead to loss of
recovery or reduced product purity, the addition of more than the
necessary quantities of cyanide reagent is not accompanied by
penalties to the same degree, except of course, the cost of the
additional reagent.
Close attention to mill feed characteristics and careful and
frequent analysis of its mineral content can result in reduction
of cyanide dosage to that actually required. In recent years,
on-line analysis techniques and reagent addition controls have
become available to minimize excess additions of reagent.
Few froth flotation process facilities in the industry have
reported treated effluent cyanide concentrations equal to or in
excess of 0.1 mg/1, and these only on an infrequent basis. Mill
6102, the largest consumer of cyanide in terms of dosage per unit
of ore feed, has been observed in the past to generate effluent
cyanide concentrations as high as 0.2 mg/1 to 6.4 mg/1. Follow-
ing installation of cyanide treatment, this facility is reporting
cyanide levels less than 0.1 mg/1. Three other flotation mills
have reported discharge concentrations in excess of 0.1 mg/1
(2122, 3121, 6101). In each case, the cyanide dosages used in
mill feed appear to be consistent with dosages reported through-
out the industry and are not unusual in that respect. Fluctua-
tions and peaking in cyanide concentrations appear to be related
to short-term overdoses of cyanide in the flotation process. Few
treated effluent measurements in the entire industry have
exceeded 0.2 mg/1 and we believe that, with close process control
and reagent addition in combination with a well designed and
operated treatment system, the 0.2 mg/1 measurement for total
cyanide can be achieved without additional treatment technology
for cyanide. For the rare case where difficulty may be
encountered or great reliability is required, treatment
technology (i.e., chemical oxidation) is available as discussed
in Section VIII and costed in Section X.
Many existing NPDES permits for ore mills contain limitations on
total cyanide. As example, in EPA Region VIII, there are nine
existing permits that limit total cyanide in the discharge from
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ore mills and these limitations vary from .02 mg/1 to .2 mg/1
daily maximum. In EPA Region X there are twelve permits for ore
mills that limit total cyanide and these limitations vary from
.01 mg/1 to .3 mg/1 daily maximum. Monitoring data for these
permits confirm that these mills are consistently within their
permit limitations on total cyanide and that the limitations can
be obtained by control of the process and the incidential removal
of cyanide as discussed below.
Incidental Cyanide Removal
Frequently, specific cyanide treatment technology is not
necessary if close process control combined with incidental
removal leads to low concentrations of total cyanide in mill
water treated effluent. This incidental removal is thought to
involve several mechanisms, including ultraviolet irradiation,
biochemical oxidation, and natural aeration. As evidence that
such mechanisms are involved, it has been noted that effluent
cyanide concentrations tend to be somewhat higher during winter
months when biological activity in the tailing pond is lower and
ultraviolet exposure is much lower due to shortened daylight
hours, less intense radiation, and ice cover on the ponds.
In addition, the association of cyanide with the depressed
minerals (i.e., pyrites) will cause a portion of the cyanide to
be removed together with the suspended solids and deposited in
the tailing ponds.
Precision and Accuracy Study
A study of the analysis of cyanide in ore mining and processing
wastewater was conducted in cooperation with the American Mining
Congress to investigate the causes of analytical interferences
observed and to determine what effect these interferences had on
the precision of the analytical method. The purpose of this
study was to evaluate the EPA-approved method and a modified
method for the determination of cyanide. The modified method
employed a lead acetate scrubber to remove sulfide compounds
produced during the reflux-distillation step. Sulfides have been
suspected of providing an interference in the colorimetric
determination of cyanide concentrations. Also, several samples
were spiked with thiocyanate to ascertain if this compound caused
interference in the cyanide analysis.
A statistical analysis of the resultant data shows no significant
difference in precision or accuracy of the two methods employed
when applied to ore mining and milling wastewaters having cyanide
concentrations in the 0.2 mg/1 to 0.4 mg/1 range. Based upon the
statistical analysis, approximately 50 percent of the overall
error of either method was attributed to intralaboratory error.
This highlights the need for an experienced analyst to perform
cyanide analyses. After considering the results of this study
and the levels achieved through dose control of reagent addition
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in the mill, EPA considered proposing an effluent limitation of
0.2 mg/1. That limitation is based on a grab sample for any one
day, and would have been subject to 100 percent error to account
for the precision and accuracy of the analytical method. (See
Section V above). Therefore, the Agency would have had to allow
an analytical measurement of up to 0.4 mg/1. However, all the
data observations in our sample were below that level.
Accordingly, the Agency is proposing to exclude cyanide from
national regulation in the ore mining category.
ADDITIONAL PARAGRAPH 8. EXCLUSIONS
Exclusion of Cyanide
Total cyanide is not regulated in BAT because the Agency cannot
quantify a reduction in total cyanide from observed
concentrations being discharged by use of technologies, known to
the Administrator, Paragraph 8(a) iii of the Settlement
Agreement.
The references to total cyanide levels of less than 0.2 mg/1
throughout this document are for informational purposes only and
are subject to the precision and accuracy of the analytical
method.as discussed here and in Section V.
Exclusion of_ Arsenic and Nickel
EPA reviewed the achievable levels calculated based on the
capabilities of the three candidate BAT treatment technologies
(secondary settling, coagulation and flocculation, and granular
media filtration). The Agency examined the necessity of
proposing specific limitations for all seven of the toxic metals
considered for regulation. Limitations on copper, lead, and zinc
are necessary since these are the metals recovered from mining
operations and concentrated in mills in this category. From a
treatability viewpoint, control of some toxic metals (arsenic and
nickel) may be achieved by limitations upon which other
pollutants are controlled. As discussed in this section, since
most of the metals are in suspended solids, reduction of arsenic
and nickel occurs in conjunction with the removal of TSS and
other toxic metals (copper, lead, and zinc).
The BAT data base for the Ore Mining and Dressing Point Source
Category was searched for instances in which arsenic and nickel
concentrations exceeded BPT limitations when copper, lead, and
zinc concentrations were also below their respective BPT limita-
tions. There was only one instance in over 300 samples in which
a nickel or arsenic concentration exceeded their BPT limitations
when BPT limitations for copper, lead, and zinc were met. The
one instance was the discharge from a sedimentation pond at
Facility 3103. The nickel concentration was 0.22 mg/1 as opposed
to the 0.20 mg/1 BPT limitation.
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The Agency concluded that the limitations on copper, lead, and
zinc would ensure adequate control of arsenic and nickel, and
under Paragraph 8(a)iii of the Settlement Agreement, arsenic and
nickel are excluded from regulation.
Exclusion of_ Asbestos
Chrysotile asbestos was detected in wastewater samples in all
subcategories and subparts within the ore mining and dressing
point source category. It was detected in 90 of 91 samples
throughout the entire industrial category.
EPA believes that the most appropriate way to regulate a toxic
pollutant is by a direct limitation on the toxic pollutant.
However, direct limitation of toxic pollutants is not always
feasible. In the case of chrysotile asbestos, there is no EPA
approved method of analysis for industrial wastewater samples.
The method of analysis presently used was developed for drinking
water samples. In addition, there are less than half a dozen
laboratories in the United States that are capable of performing
the analysis by this method.
Chrysotile asbestos is known to be present in many ore deposits
throughout the country (Reference 6). As ore is mined and
subsequently milled, it is subjected to a variety of crushing and
size reduction operations. As a result, smaller solids are
formed, the chrysotile asbestiform fibers are liberated as the
small solids are made, and end up in mine drainage and mill
process water.
The possibility of the chrysotile asbestos fibers being present
in waste streams for the same reasons and in the same relative
proportions as the solids, led to the examination of the EGD
sampling data in an attempt to establish a relationship between
chrysotile asbestos and TSS.
Review of analyses for asbestos and TSS in samples of untreated
and treated wastewater shows that as TSS is reduced by treatment,
observed asbestos concentrations are also reduced.
Intake water samples (from 26 industrial categories) and POTW
effluents were reviewed to get an indication of background levels
for chrysotile asbestos. The values of chrysotile asbestos
ranged from 3.5 x 1 O4 (detection limit) to 1.63 x 108 with a mean
value of 1.1 x 1O7 fibers per liter. The treated waste stream
sample values for chrysotile asbestos in the ore data ranged from
10* (detection limit) to 108 fibers per liter.
The Agency has determined that when TSS is reduced to the BPT
effluent limitations for ore mining and dressing, observed
chrysotile asbestos levels are reduced to or below background
levels. Therefore, EPA is excluding chrysotile asbestos from
regulation since it is effectively controlled by technologies
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upon which TSS limitations are based, Paragraph 8(a)iii of the
Settlement Agreement.
Exclusion of_ Pollutants Detected ir\_ a Single Source and Uniquely
Related to That Source
There are 19 operating uranium mills in the United States, 18 of
which now achieve zero discharge of process wastewater. There
are no uranium mills that commingle process wastewater with mine
drainage and it is anticipated that none of these zero discharge
mills would elect to treat and discharge at the BPT limitations
because of the expense to install technology required, i.e., ion
exchange, ammonia stripping, lime precipitation, barium chloride
coprecipitation, and settling.
EPA is excluding uranium mills from BAT, since there is only one
discharging facility and it is believed that none of the other
existing facilities will commingle mine drainage and mill process
wastewater. Uranium mills are not regulated in BAT because the
pollutants found in the discharge are uniquely related to this
single source, Paragraph 8(a)iii of the Settlement Agreement.
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SECTION XI
BEST CONVENTIONAL POLLUTANT CONTROL TECHNOLOGY
Section 301(b)(2)(E) of The Act requires that there be achieved,
not later than July 1, 1984, effluent limitations for categories
and classes of point sources, other than publicly-owned treatment
works, that require the application of the best conventional
pollutant control technology (BCT) for control of conventional
pollutants as identified in Section 304(a)(4). The pollutants
that have been defined as conventional by the Agency, at this
time, are biochemical oxygen demand, suspended solids, fecal
coliform, oil and grease, and pH.
BCT is not an additional limitation; rather, it replaces BPT for
the control of conventional pollutants. BCT must be evaluated
for cost effectiveness and a comparison made between the cost and
level of reduction of conventional pollutants from the discharge
of publicly owned treatment works (POTW) and the cost and level
of reduction of such pollutants from a class or category of
industrial sources.
The technologies considered for treatment of conventional
pollutants are the same as those considered for treatment of
toxic pollutants. As discussed in Section X, the Agency has
determined that the BAT limitations for toxics are equivalent to
BPT limitations for the ore mining and dressing industry. BCT
limitations for the conventional pollutants, TSS, and pH are also
proposed at BPT levels. Accordingly, by definition, BCT for this
industry meets any BCT cost test because there is no incremental
cost to remove conventional pollutants beyond BPT.
Summary of Best Conventional Technology
Daily 30-Day
Maximum Average
Pollutant (mg/1) (mq/1)
TSS 30 20
pH Within the range of
6 to 9
The Agency is currently developing a new BCT cost methodology.
It is possible, though unlikely, that a treatment technology more
stringent than BPT will provide additional removal of
conventional pollutants and pass the new cost test, when
developed. In that event, the Agency will propose new BCT
limitations.
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SECTION XII
NEW SOURCE PERFORMANCE STANDARDS (NSPS)
The basis for new source performance standards (NSPS) under
Section 306 of the Act is through application of the best avail-
able demonstrated technology. New facilities have the opportu-
nity to implement the best and most efficient ore mining and
milling processes and wastewater technologies. Congress, there-
fore, directed EPA to consider the best demonstrated process
changes and end-of-pipe treatment technologies capable of
reducing pollution to the maximum extent feasible.
GENERAL PROVISIONS
Several items of discussion apply to options in more than one
subcategory. To avoid repetition, these items are discussed
here.
Relief From N_£ Discharge Requirement
For new sources that must achieve no discharge of process
wastewater, the Agency provides relief upon the occurrence of a
10-year, 24-hour precipitation event or snow melt of equivalent
volume. Excess water resulting from the occurrence shall not be
subject to effluent limitations.
Relief from no discharge of process wastewater is also granted
for those facilities located in net precipitation areas and is
the same relief granted to BAT limitations requiring no
discharge.
Relief From Effluent Limitations for Those Facilities Permitted
to Discharge
The relief is exactly the same as that granted for existing
sources under BAT. Excess water resulting from precipitation
from a facility designed, constructed, and maintained to contain
or treat the maximum volume of process wastewater discharged
during any 24-hour period, including the volume that would result
from a 10-year, 24-hour precipitation event or snow melt of
equivalent volume, shall not be subject to the limitations set
forth in 40 CFR 440.
Commingling Provisions
For new sources that combine for treatment waste streams from
various sources, the quantity and quality of each pollutant or
pollutant property in the combined discharge that is subject to
effluent limitations shall not exceed the quantity and quality of
each pollutant or pollutant property that would have been
discharged had each waste stream been treated separately. The
discharge flow from a combined discharge shall not exceed the
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volume that would have been discharged had each waste stream been
treated separately.
NSPS OPTIONS CONSIDERED
The Agency considered the following NSPS options:
Option One. Require achievement of performance standards in each
subcategory based on the same technology as BAT (NSPS = BAT).
Option Two. Require standards based on a complete water recycle
system (NSPS = zero discharge).
NSPS SELECTION AND DECISION CRITERIA
EPA has selected performance standards based on the same
technology as BAT for all facilities in the ore mining and
dressing point source category, except those facilities using
froth flotation in the copper, lead, zinc, gold, silver,
platinum, and molybdenum subcategory and mills in the uranium
subcategory.
Subcategories and Subparts Under Option ]_
Option 1 (NSPS = BAT) has been selected for iron ore mills in the
Mesabi range; copper, lead, zinc, silver, gold, platinum, and
molybdenum mills that use leaching to recover copper and the
cyanidation process for the recovery of gold; and mercury mills
since BAT specifies zero discharge. Option 1 (NSPS = BAT) has
also been selected for iron ore mine drainage, iron ore mills,
aluminum mine drainage, copper, lead, zinc, gold, silver,
platinum, and molybdenum mine drainage, titanium mine drainage,
dredges and mills, and mercury mine drainage. The concentration
levels of toxic metals found in new sources in these
Subcategories and subparts are expected to be similar to existing
sources. Following the implementation of BAT, toxic metals will
be found at or near detection levels or at concentrations below
the practical limits of additional technology. Further reduction
of these pollutants can not be technically or economically
justified.
Subcategories and Subparts Under Option 2 (NSPS = zero discharge)
EPA is proposing that new source froth flotation mills achieve
zero discharge of process wastewater. EPA considered zero
discharge based on recycle for existing copper, lead, zinc, gold,
silver, platinum, and molybdenum mills using froth flotation, but
rejected it because of the extensive retrofit required at some
existing facilities, the cost of retrofitting, and the possible
changes required in the process. This concern does not apply to
new sources. Zero discharge is a demonstrated technology at 46
of the 90 froth flotation mills for which EPA has data (see
wastewater discharges as summarized in Tables IX-2 through IX-10)
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and meets the definition of standard of performance permitting
zero discharge of pollutants. New sources have the option to
recycle because the metallurgical process can be adjusted and
designed to recycle process wastewater before the actual
construction of the new source. While reagent buildup has been
mentioned by industry as a potential problem in extractive
metallurgy, no evidence has been submitted to validate ths
assertion.
There are new sources anticipated in copper, lead, zinc, gold,
silver, and molybdenum mining. Standards applied to these new
source waste streams should reflect the best treatment levels
achievable by the froth flotation segment of the industry.
A study of existing froth flotation mills reveals that a large
percentage of these facilities are effectively achieving 100
percent recycle of mill water. Many of the facilities practicing
100 percent recycle are located in arid regions, but some
facilities are located in humid regions. A summary of some
existing facilities follows.
Copper Ore
Of the 35 known froth flotation copper mills in the United
States, 31 achieve zero discharge of process wastewater.
Lead/Zinc Ores
Five of the 27 active froth flotation mills in the lead/zinc
subcategory achieve zero discharge.
Gold
Four of the five primary gold facilities employing froth
flotation techniques discharge process wastewater.
Silver
Three of the four known primary silver facilities which use froth
flotation methods are achieving zero discharge of process
wastewater.
Molybdenum
Of the three molybdenum operations employing the froth flotation
process, one facility achieves zero discharge of recycle.
In Section IX (Table IX-3 through Table IX-10) cost comparisons
for treatment technologies are made for existing sources. It is
believed that new source mills will have similar mill capacity,
water use per ton of ore, and pollutants in the raw wastewater as
existing mills. Therefore, costs of technology for new sources
will approach those costs determined for existing sources. The
527
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cost of 100% recycle for mills is generally less than the cost
determined for pH adjustment of the wastewater (lime addition).
Assuming the same settling pond, or tailings pond, would be
required for recycle as for pH adjustment and settling, the cost
to recycle at a new source is approximately the same, or less,
than the cost to implement the technology (pH adjustment and
settle) upon which BPT and BAT effluent limitations are based.
EPA is proposing that new source uranium mills achieve zero
discharge of process wastewater. For this subpart, EPA
considered zero discharge for BAT based on total impoundment and
evaporation, or recycle and reuse of the mill process water, or a
combination of these technologies. Because the pollutants
detected in the current discharge from this subpart are uniquely
related to one point source, the single mill discharging, the
uranium mill subpart is excluded from BAT under Paragraph 8
authority of the Settlement Agreement (see Section X).
However, the Agency believes that for new sources a standard of
performance must be proposed. Otherwise, additional discharges
(new sources) could occur that obviously would not be unique to
one source. New source uranium mills are anticipated by the
Agency. New mill capacity is anticipated to replace existing
mills and to maintain the current production of uranium oxide as
lower grade ore must be mined. Also, an increase in demand for
uranium oxide is anticipated that will require new mills.
Uranium mills can achieve zero discharge as indicated by the fact
that 18 of 19 existing mills currently achieve no discharge.
As discussed in Section X, ammonia stripping, lime precipitation,
barrium chloride co-precipitation or ion exchange, and settling
are the identified technologies for uranium mills to meet BPT
limitations on metals, radium 226, ammonia and TSS. The cost to
implement the technologies to meet the BPT limitations for a new
uranium mill is more than the cost to implement recycle or
evaporation ponds (or the combination of the two) to meet a no
discharge requirement.
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SECTION XIII
PRETREATMENT STANDARDS
Section 307(b) of the Act requires EPA to promulgate pretreatment
standards for both existing sources (PSES) and new sources (PSNS)
of pollution which discharge their wastes into publicly owned
treatment works (POTWs). These pretreatment standards are
designed to prevent the discharge of pollutants which pass
through, interfere with, or are otherwise incompatible with the
operation of POTWs. In addition, the Clean Water Act of 1977
adds a new dimension of these standards by requiring pretreatment
of pollutants, such as heavy metals, that limit POTW sludge
management alternatives. The legislative history of the Act
indicates that pretreatment standards are to be technology based
and, with respect to toxic pollutants, analogous to BAT. The
Agency has promulgated general pretreatment regulations which
establish a framework for the implementation of these statutory
requirements (see 43 FR 27736, 16 June 1978).
EPA is not proposing pretreatment standards for existing sources
(PSES) or new sources (PSNS) in the ore mining and dressing point
source category at this time nor does it intend to promulgate
such standards in the future since there are no known or
anticipated discharges to publicly owned treatment works (POTWs).
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SECTION XIV
ACKNOWLEDGEMENTS
This document was drafted by Radian Corporation, McLean, Virginia
under the direction of Mr. Baldwin M. Jarrett, Project Officer,
Energy and Mining Branch, Effluent Guidelines Division, EPA.
Direction and assistance were also provided by Mr. William A.
Telliard, Chief of the Energy and Mining Branch and Mr. Ron
Kirby, Project Monitor. The insights and review provided by Mr.
Barry Neuman, formerly of the Office of General Counsel, are also
expressly appreciated. Much of the input for this document was
provided by Radian's subcontractors Frontier Technical
Associates, Buffalo, New York, and Hydrotechnic Corporation, New
York, New York.
The following divisions of the Environmental Protection Agency
(EPA) contributed to the development of this document: All
regional offices/ Industrial Environmental Research Laboratory
Cincinnati, Ohio; Office of Research and Development; Office of
General Counsel; Office of Planning and Evaluation; Monitoring
and Data Support; Criteria and Standards Division; Office of
Quality Review; and Office of Analysis and Evaluations.
Appreciation is extended to the following trade associations and
mining companies for assistance and cooperation during the course
of this program:
Aluminum Association
American Iron Ore Association
American Mining Congress
Aluminum Company of America
Amax Lead Company of Missouri
American Exploration and Mining Company
American Smelting and Refining Company
Anaconda Copper Company
Atlas Corporation
Bethlehem Mines Corporation
Brush Wellman Incorporated
Bunker Hill Company
Carl in Gold Mining Company
Cities Service Company
Cleveland-Cliffs Iron Company
Climax Molybdenum Company
Cominco American, Inc.
Continental Materials Corporation
Copper Range Company
Curtis Nevada Mines, Inc.
Cyprus-Bagdad Copper Corporation
Eagle Pitcher Industries, Inc.
E. I. DuPont de Nemours and Company, Inc.
Erie Mining Company
Goodnews Bay Mining Company
531
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Hanna Mining Company
Heela Mining Company
Homestake Mining Company
Idarado Mining Company
Inspiration Consolidated Copper Company
Jones and Laughlin Steel Corporation
Kennecott Copper Corporation
Kerramerican, Inc.
Kerr McGee Corporation
Knob Hill Mines, Inc.
Lead and Zinc Institute
Magma Copper Company
Marquette Iron Mining Company
Molybdenum Corporation of America
National Lead Industries, Inc.
New Jersey Zinc Company
Oat Hill Mining Company
Oglebay-Norton Company - Eveleth Taconite
Phelps Dodge Corporation
Pickands Mather and Company - Erie Mining Company
Ranchers Exploration and Development Corporation
Rawhide Mining Company
Reynolds Mining Corporation
Standard Metals Corporation
St. Joe Minerals Company
Sunshine Mining Company
Titanium Enterprises
Union Carbide Corporation
United Nuclear Corporation
U.S. Antimony Corporation
U.S. Steel Corporation
White Pine Copper Company
The initial draft of this document containing most of the
information and data on which EPA relies was developed and
written by Calspan Corporation, Buffalo, New York. Copies of
this draft were distributed to Federal and State agencies, trade
associations, conservation organizations, industry, and
interested citizens. We wish to thank the following who
responded with written comments: American Mining Congress;
Bunker Hill Company; Natural Resources Defense Council, Inc.;
Prather, Seeger, Doolittle, and Farmer; St. Joe Minerals
Corporation, Trustees for Alaska; U.S. Department of Interior -
Bureau of Mines; U.S. Department of Labor; USEPA - Environmental
Research Laboratory (Athens, Georgia); Walter C. McCrone
Associates, Inc.; White Pine Copper Company. An effort has been
made in this document to address those comments which pointed to
deficiencies in the record and data contained in the draft.
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SECTION XV
REFERENCES
SECTION III
1. Dennis, W. H., Extractive Metallurgy Principles and
Application, Sir Issac Pitman & Sons, Ltd., London, 1965.
2. U.S. Bureau of Mines, A Dictionary of Mining, Mineral and
Related Terms. Compiled and edited by P. W. Thrush and the USBM
Staff. Washington, D.C.: U.S. Department of Interior (1968).
3. Cummins, A. B. and I. A. Given, Mining Engineering 'Handbook,
Volumes I and II. Littleton, Colorado; Society of Mining
Engineers (1973) .
4. "Iron Ore 1976," American Iron Ore Association, Cleveland,
Ohio, 1976.
5. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1978/1979.
6. Personal communications from Commodity Specialist, Bureau of
Mines, U.S. Department of the Interior, to M. A. Wilkinson,
Calspan Corporation, January 1978.
7. "Metals Statistics," American Metal Market, Fairchild
Publications, Inc., New York, 1974.
8. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1975.
9. The Wall Street Journal, 16 September 1981, p. 46.
10. "A Study of Waste Generation, Treatment and Disposal in the
Metals Mining Industry," Midwest Research Institute, October
1976.
11. "Metal-Nonmetal Mine File Reference," Mining and Engineering
Safety Administration, Washington, 9 March 1977.
12. 1979/E/MJ International Directory o_f Mining and Mineral
Processing Operations, Engineering and Mining Journal, New York,
1979.
13. Engineering and Mining Journal, McGraw-Hill, February 1980,
Vol. 181, No. 2, p. 25.
14. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1974.
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15. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 556, 1963.
16. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 650, 1970.
17. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 667, 1975.
1^. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1977.
19. Personal communications from J. Viellenave, Earth Sciences,
Inc., to M. A. Wilkinson, Calspan Corporation, October 1976.
20. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1973.
21. Personal communication from Linda Carrico, commodity
specialist U.S. Department of the Interior, to D. M. Harty,
Frontier Technical Associates, April 1981.
22. Gordon, E., "Uranium—Rising Prices Continue to Spur
Development," Engineering and Mining Journal, March 1977.
23. White, G., Jr., "Uranium: Prices Steady at High Level in
'77," Engineering and Mining Journal, March 1978.
24. Merrit, R. C., "Extractive Metallurgy of Uranium," Colorado
School of Mines Research Institute, Inc., Colorado, 1971.
25. Davis, J. F., "U.S. Uranium Industry Continues Active
Development Despite Nuclear Uncertainties," Engineering and
Mining Journal, August 1977.
26. U.S. Nuclear Regulatory Commission, "Draft Generic Environ-
mental Impact Statement on Uranium Milling," Report No. NUREG-
0511, April 1979.
27. Personal communication from Scott F. Sibley, commodity
specialist U.S. Department of the Interior, to D. M. Harty,
Frontier Technical Associates, May 1981.
SECTION V
1. Bainbridge, Kent, "Evaluation of Wastewater Treatment
Practices Employed at Alaskan Gold Placer Mining Operations,"
Calspan Report No. 6332-M-2, Prepared for U.S. Environmental
Protection Agency - Effluent Guidelines Division, Washington,
D.C., 17 July 1979.
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2. Harty, David M. and P. Michael Terlecky, "Titanium Sand
Dredging Wastewater Treatment Practices," Frontier Technical
Associates, Inc., Buffalo, New York, Report No. 1804-1, Prepared
for U.S. Environmental Protection Agency - Effluent Guidelines
Division, Washington, D.C., Revised 20 October 1980.
3. Hayes, B. J., R. M. Mann, and J. I. Steinmetz, "Cyanide
Methods Evaluation: A Modified Method for the Analysis of Total
Cyanide in Ore Processing Effluents" Radian Corporation DCN No.
80-210-002-14-09, Prepared for U.S. Environmental Protection
Agency - Effluent Guidelines Division, Washington, D.C., 11
August 1980.
SECTION VI
1. "Froth Flotation in 1975," Advance Summary of Mineral
Industry Surveys, Bureau of Mines, U.S. Department of the
Interior, Washington, 1976.
2. Hawley, J. R., "The Use, Characteristics and Toxicity of
Mine-Mill Reagents in the Province of Ontario," Ontario (Canada)
Ministry of the Environment, Ottawa, 1972.
3. Mining Chemicals Handbook, Mineral Dressing Notes No. 26,
American Cyanamid Company, Wayne, NJ, 1976.
4. Sharp, F. H., "Lead-Zinc-Copper Separation and Current
Practices at the Magmont Mill," Engineering and Mining Journal,
July 1973.
5. Personal communication from Chemist and Concentrator General
Foreman, Lead/Zinc Mine/Mill 3103, to P. H. Werthman, Calspan
Corporation, 1978.
SECTION VII
1. "Condensed Chemical Dictionary," P. Hawley, Van Norstrand,
Reinhold, New York, New York, 1971.
2. Rawlings, G. D., and M. Samfield, Environmental Science and
Technology, Vol. 13, No. 2, February 1974.
3. Development Document for BAT Effluent Limitations Guidelines
and New Source Performance Standards for the Textiles Point
Source Category, U.S. EPA, EGD, 1979.
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4. "Sampling and Analysis Procedures for Screening of Industrial
Effluents for Priority Pollutants," U.S. Environmental
Protection Agency, Environmental Monitoring and Support
Laboratory, Cincinnati, Ohio, March 1977, Revised April 1977.
5. "Seminary for Analytical Methods for Priority Pollutants,"
U.S. Environmental Protection Agency, Office of Water Programs,
Savannah, Georgia, 23- 24 May 1978.
6. "Antimony Removal Technology for Mining Industry Waste--
waters," Draft Report Prepared for U.S. Environmental Protection
Agency by Hittman Associates, Columbia, MD., under Contract 68-
03-1566, HIT-C185/200-78-739D, July 1978.
7. Calspan Corporation, "Heavy Metal Pollution from Spillage at
Ore Smelters and Mills," Calspan Report No. ND-5187 M-l (Rev),
EPA Contract No. 68-01-0726, Prepared for U.S. Environmental
Protection Agency, Office of Research and Development, Industrial
Environmental Research Laboratory, Cincinnati, Ohio, 15 July
1977.
8. Patterson, J. W., Wastewater Treatment Technology, Ann Arbor
Science Publishers, Inc., Ann Arbor, Michigan, 1977.
9. "Development Document for Effluent Limitations Guidelines and
New Source Performance Standards for the Ore Mining and Dressing
Point Source Category," U.S. Environmental Protection Agency, EPA
440/178/061-e, PB-286 521, July 1978.
10. Terlecky, P. M., editor, "Draft Development Document for the
Miscellaneous Nonferrous Metals Segment of the Nonferrous Metals
Point Source Category," EPA-440/1-76-067, March 1977.
SECTION VIII
1. Mining Chemicals Handbook, Mineral Dressing Notes No. 26,
American Cyanamid Company, Wayne, NJ, 1976.
2. Personal communication from Chemist and Concentrator General
Foreman, Lead/Zinc Mine/Mill 3103, to P. H. Werthman, Calspan
Corporation, 1978.
3. Personal communication from Water Quality Project Engineer,
Copper Mine/Mill 2122, to K. L. Bainbridge, Calspan Corporation,
1978.
4. Personal communication from Mill Superintendent, Lead/Zinc
Mine/Mill 3101, to K. L. Bainbridge, Calspan Corporation, 1978.
536
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5. Miller, J., "Pyrite Depression by Reduction of Solution
Oxidation Potential," Department of Mineral Engineering,
University of Utah, Salt Lake City, 1970.
6. "Alternative for Sodium Cyanide for Flotation .Control
(T2008)," Report Prepared for U.S. Environmental Protection
Agency, Cincinnati, Ohio, Battelle—Columbus (Ohio) Laboratories,
1979.
7. "Treatability of and Alternatives for Sodium Cyanide for
Flotation Control," Report for U.S. Environmental Protection
Agency, Cincinnati, Ohio, by Battelle - Columbus (Ohio)
Laboratories, 31 January 1980.
8. Culp, R. L. and G. L. Culp, Advanced Wastewater Treatment,
Van Nostrand Reinhold Book, Co., New York, 1971.
9. "Process Design Manual for Suspended Solids Removal," U.S.
Environmental Protection Agency, Washington, EPA 625/1-75-003a,
January 1975.
10. Bernardin, F. E., "Cyanide Detoxification Using Adsorption
and Catalytic Oxidation on Granular Activated Carbon," Journal of
Water Pollution Control Federation, Vol. 45, No. 2, February
1973.
11. Bucksteeg, W. and H. Thiele, "Method for the Detoxification
of Wastewater Containing Cyanide," German Patent 1,140,963,
January 1969.
12. Kuhn, R., "Process for Detoxification of Cyanide Containing
Aqueous Solutions," U.S. Patent 3,586,623, June 1971.
13. Willis, G. M. and J. T. Woodcock, "Chemistry of Cyanidation
II: Complex Cyanides of Zinc and Copper," Proc. Austral. Inst.
Min. Metall., No. 158-159, 1950, 00. 465-488.
14. Woodcock, J. T. and M. H. Jones, "Oxygen Concentrations,
Redox Potentials, Xanthate Residuals, and Other Parameters in
Flotation Plant Pulps," Proceedings of the Ninth Commonwealth
Mining and Metallurgial Congress, The Institute of Mining and
Metallurgy, London (England), 1969.
15. Letter from Vice President, Metallurgy, Copper Mine/Mill
2138 and Lead/Zinc Mine/Mills 3120 and 3121, to R.B. Schaffer,
Director, Effluent Guidelines Division (WH-552), U.S.
Environmental Protection Agency, 9 May 1977.
16. Written comments from Owner, Copper Mine/Mills 2101, 2104,
2116, and 2122 and Lead/Zinc Mine/Mill 3142, to U.S.
537
-------
Environmental Protection Agency, Washington, relative to interim
final effluent limitations for ore mining and dressing industry,
7 January 1976.
17. Eccles, A. G., "Cyanide Destruction at Western Mines Myra
Falls Operation," Paper presented at CMP Annual Meeting, 25
January 1977.
18. Eiring, L. V., Uchenye Zapiski Universities, Erevan, No. 2,
1967.
19. Eiring, L. V., "Kinetics and Mechanism of Ozone Oxidation of
Cyanide-Containing Wastewater," Soviet Journal of Non-Ferrous
Metals, Vol. 55, No. 106, 1969, pp. 81-83.
20. Chamberlin, N. S. and H. B. Snyder, Jr., "Treatment of
Cyanide and Chromium Wastes," Proceedings of_ Regional Conference
on Industrial Health, Houston, Texas, 27-29 September 1951.
21. Chamberlin, N. S. and H. B. Snyder, Jr., "Technology of
Treating Plating Wastes," Paper presented at Tenth Industrial
Waste Conference, Purdue University, West Lafayette, Ind., 9-11
May 1955.
22. Patterson, J. W., Wastewater Treatment Technology, Ann Arbor
Science Publishers, Inc., Ann Arbor, Mich., 1977, Chapter 9,
"Treatment Technology for Cyanide."
23. Goldstein, M., "Economics of Treating Cyanide Wastes,"
Pollution Engineering, March 1976.
24. Bird, A. 0., "The Destruction and Detoxification of Cyanide
Wastes," Chemical Engineer, University of Birmingham (England),
1976, pp. 12-21 .
25. Bollyky, L. J., "Ozone Treatment of Cyanide Plating Wastes,"
Paper presented at First International Symposium on Ozone Water
and Wastewater Treatment, 1973.
26. Personal communication from R. Mankes, Telecommunications
Industries, Inc., to R. C. Lockemer, Calspan Corporation,
November 1977.
27. Fiedman, I. D. et al., "Removal of Toxic Cyanides from
Wastewaters of Gold Extracting Mills," Sb. Mosk. Inst. Stali
Splavov (USSR), No. 53, 1969, pp. 106-116 and Chemical Abstracts,
Vol. 72, No. 12, 1970, pp. 272-273.
538
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28. Boriesnko, A. P., "Detoxification of Wastewaters from the
Zolotushinskii Beneficiation Mills," Tsvet. Metal. (USSR), Vol.
42, No. 6, 1969, pp. 16-18.
29. Eiring, L. V., "Processing of Wastewaters from Gold
Extraction Plants," Vodosnabzh. i_ Sanit. Tekln (USSR), (2), 4-5,
1965.
30. Garrison, R. L., C. E. Mauk, and W. Prengle, Jr., "Cyanide
Disposal by Ozone Oxidation," Air Force Weapons Laboratory,
Kirtland Air Force Base, NM, AFWL-TR-73-212, February 1974.
31. Lanouette, K. H., "Treatment of Phenolic Wastes," Chemical
Engineering, Deskbook Issue, 17 October 1977.
32. Patterson, J. W., Wastewater Treatment Technology, Ann Arbor
Science Publishers, Inc., Ann Arbor, Mich., 1975.
33. Rosfjord, R. E., R. B. Trattner, and P. N. Cheremisinoff,
"Phenols: A Water Pollution Control Assessment," Water and
Sewage Works, March 1976.
34. Hodgson, E. W., Jr., and P. M. Terlecky, Jr. (Ed.), "Adden-
dum to Development Document for Effluent Limitations Guidelines
and New Source Performance Standards for Major Inorganic Products
Segment of Inorganic Chemicals Manufacturing Point Source
Category," Prepared for Effluent Guidelines Division, U.S.
Environmental Protection Agency, Washington, by Calspan
Corporation, Buffalo, NY, ND-5782-M-72, June 1978.
35. Larsen, H. P., "Chemical Treatment of Metal-Bearing Mine
Drainage," Journal of Water Pollution Control Federation, Vol.
45, No. 8, 1973, pp. 1682-1695.
36. Shimoiizuka, J., "Recovery of Xanthates from Cadmium
Xanthate," Nippon Kogyokaishi, 88, 1972, pp. 539-543.
37. Rosehard, R. and J. Lee, "Effective Methods of Arsenic
Removal from Gold Mine Wastes," Canadian Mining Journal, June
1972, pp. 53-57.
38. Curry, N. A., "Philosophy and Methodology of Metallic Waste
Treatment," Paper presented at 27th Industrial Waste Conference,
Purdue University, West Lafayette, Ind., 1972.
39. Larsen, H. P. and L. W. Ross, "Two Stage Process Chemically
Treats Mine Drainage to Remove Dissolved Metals," Engineering and
Mining Journal, February 1976.
40. Terlecky, P. M., Jr., M. A. Bronstein, and D. W. Goupil,
"Asbestos in the Ore and Dressing Identification Analysis,
Treatment Technology, Health Aspects, and Field Sampling
Results," Prepared for Effluent Guidelines Division, U.S.
539
-------
Environmental Protection Agency, Washington, by Calspan
Corporation, Buffalo, NY, ND-5782-M-121, (1979).
41. Logsdon, G. S. and J. M. Symons, "Removal of Asbestiform
Fibers by Water Filtration, Journal of_ American Water Works
Association, September 1977, pp. 499-506.
42. "Direct Filtration of Lake Superior Water from Asbestiform
Fiber Removal," Prepared for U.S. Environmental Protection Agency
EPA-670/2-75-050, by Black and Veatch Consulting Engineers, 1975.
43. Robinson, J. H. et al., "Direct Filtration of Lake Superior
Water for Asbestiform - Solids Removal," Journal of_ American
Water Works' Association, October 1976, pp. 531-539.
44. Lawrence, J. and H. W. Zimmerman, "Asbestos in Water: Mining
and Processing Effluent Treatment," Journal of_ Water Pollution
Control Federation, January 1977, pp. 156-160.
45. Patton, J. L., "Unusual Water Treatment Plant Licks Asbestos
Fiber Problem," Water and Wastes Engineering, 50, November 1977,
pp. 41-44. 46. Yourt, R. G., "Radiological Control of Uranium
Mine and Mill Wastes," Proceedings of the 13th Conference on
Industrial Waste, Ontario (Canada), 1966, pp. 107-120.
47. Beverly, R. G., "Unique Disposal Methods are Required for
Uranium Mill Waste," Mining Engineering (Transaction, AIME) 20,
1968, pp. 52-56.
48. Felman, M. H., "Removal of Radium from Acid Mill Effluents,"
WIN-125, 1961.
49. Ryan, R. K. and P. G. Alfredson, "Liquid Wastes from Mining
and Milling of Uranium Ores: A Laboratory Study of Treatment
Methods," ISBN 0-642-99752-4, AAEC/E394, Australian Heights,
October 1976.
50. "The Control of Radium and Thorium in the Uranium Milling
Industry," United States Atomic Energy Commission, WIN-112, 1960.
51. Arnold, W. D. and D. J. Grouse, "Radium Removal from Uranium
Mill Effluents with Inorganic Ion Exchangers," Ind. Egn. Chem.
Process Des. Dev., Vol. 4, No. 3, 1965, pp. 333-337.
52. Clark, J. W., W. Viessman, Jr., M. J. Hammer, Water Supply
and Pollution Control, Third Edition, Harper and Row Publishers,
New York, N. Y., 1977.
540
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53. Metcalf and Eddy, Wastewater Engineering: Treatment,
Disposal, Reuse, Second Edition, McGraw-Hill Book Company, New
York, N. Y., 1979.
54. "Mine Wastewater Pilot Treatment Project—Metal Removal
Phase," Montreal (Quebec, Canada), Engineering Company, May 1975.
55. Campbell, H. and B. P. LeClair, "Dewatering Base Metal Mine
Drainage Sludge," Proceedings of 1Oth Canadian Symposium on Water
Pollution Research, 1975.
56. Huck, P. M. and B. P. LeClair, "Operational Experience with
a Base Metal Mine Drainage Pilot Plant," EPA 4-WP-74-8, September
1974; Also presented as paper at 29th Industrial Waste
Conference, Purdue University, West Lafayette, Ind., May 1974.
57. Huck, P. M. and B. P. LeClair, "Treatment of Base Metal Mine
Drainage," Paper presented at 30th Industrial Waste Conference,
Purdue University, West Lafayette, Ind., 1975.
58. Huck, P. M. and B. P. LeClair, "Polymer Selection and Dosage
Determination Methodology for Acid Mine Drainage and Tailings
Pond Overflows," Paper presented at Symposium on Flocculation and
Stabilization of Solids in Aqueous and Non-Aqueous Media, Toronto
(Ontario, Canada), November 1974.
59. "Pilot-Scale Wastewater Treatability Studies Conducted at
Various Facilities in the Ore Mining and Milling Industry,"
Prepared for Effluent Guidelines Division, U.S. Environmental
Protection Agency, Washington, by Calspan Corporation, Buffalo,
N. Y., 6332-M-2, in publication (1979).
60. Dean, J. G., F. L. Bosque and K. H. Lanouette, "Removing
Heavy Metals from Wastewater," Env. Sci. and Tech., Vol. 6, No.
6, 19721, pp. 518-522.
61. Nilsson, R., "Removal of Metals by Chemical Treatment of
Municipal Wastewater," Water Research, Vol. 5, 1971, pp. 51-60.
62. Lanouette, K. H. and E. G. Paulson, "Treatment of Heavy
Metals in Wastewater," Pollution Engineering, October 1976, pp.
55-57.
63. Hem, J. D., "Reactions of Metal Ions at Surfaces of Hydrous
Iron Oxide," Geochimica et Cosmochimica Acta, Vol. 41, 1977, pp.
527-538.
64. Michalovic, J. G., J. G. Fisher and D. H. Bock, "Suggested
Method for Vanadate Removal from Mill Effluent," J_._ Environ.
Sci. Health, Vol. A12, Nos. 1 and 2, 1977, pp. 21-27.
541
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65. Kunz, R. G., J. F. Giannelli and H. D. Stensel, "Vanadium
Removal from Industrial Wastewaters," Jour. Water Poll. Control
Fed., Vol. 48, No. 6, 1976, pp. 762-770.
66. LeGendre, G. R. and D. D. Runnells, "Removal of Dissolved
Molybdenum from Wastewaters by Precipitates of Ferric Iron," Env.
Sci. and Tech., Vol. 9, No. 8, 1975, pp. 755-759.
67. Gott, R. D., "A Discussion and Review of the Development of-
Wastewater Treatment at the Climax Mine, Climax, Colorado," Paper
presented at 1977 American Mining Congress, San Francisco,
California.
68. Federal Register, Vol. 43, No. 133, 11 July 1978, p. 29773.
69. Bainbridge, K., "Evaluation of Wastewater Treatment
Practices Employed at Alaskan Gold Placer Mining Operations,"
Calspan Report No. 6332-M-2 prepared for U.S. Environmental
Protection Agency Effluent Guidelines Division, Washington, D.C.
70. Jackson, B. et al., "Environmental Study on Uranium Mills—
Part I," Draft Final Report Prepared for Effluent Guidelines
Division, U.S. Environmental Protection Agency, Washington, by
TRW, Inc., under Contract 68-03-2560, December 1978.
71. Letter from D. J. Kuhn, Secured Landfill Contractors, Inc.,
Tonawanda, NY, to P. M. Terlecky, Jr., Frontier Technical
Associates, Inc., Buffalo, NY, 12 February 1979.
SECTION IX
1. Engineering News Record, Volume 203, Number 24, 13 December
1979.
2. Robert Snow Means Company, Building Construction Cost Data,
Robert Snow Means Co., Duxbury, Mass., 1980.
3. Dodge Building Cost Services, Construction Systems Costs,
McGraw Hill, 1979.
4. Harty, David M. and P. Michael Terlecky, "Characterization of
Wastewater and Solid Wastes Generated in Selected Ore Mining
Subcategories (Sb, Hg, Al, V, W, Ni, Ti)," Contract No. 68-01-
5163, Frontier Technical Associates Report No. 2804-1, 24 August
1981, Prepared for U.S. Environmental Protection Agency,
Washington, D.C.
5. PEDCo Environmental, Inc., "Evaluation of Best Management
Practices for Mining Solid Waste Storage, Disposal and Treatment
Presurvey Study," February 1981, Contract No. 68-03-2900,
542
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Prepared for U.S. Environmental Protection Agency, Office of
Research and Development, Cincinnati, Ohio.
6. Environmental Protection Agency, "Mining Industry Solid
Waste, an Interim Report," Office of Solid Waste, February 1981.
SECTION X
1. "Development Document for BPT Effluent Limitations Guidelines
and New Source Performance Standards for the Ore Mining and
Dressing Industry," U.S. Environmental Protection Agency, 1975.
2. "Process Design Manual for Suspended Solids Removal," U.S.
Environmental Protection Agency, Washington, EPA 625/1-75-003a,
January 1975.
3. Personal communication from R. Mankes, Telecommunications
Industries, Inc., to R. C. Lockemer, Calspan Corporation,
November 1977.
4. Bainbridge, K., "Evaluation of Wastewater Treatment Practices
Employed at Alaskan Gold Placer Mining Operations," Calspan
Report No. 6332-M-2 prepared for U.S. Environmental Protection
Agency Effluent Guidelines Division, Washington, D.C.
5. "Dana's Manual of Mineralogy," 17th Edition, 1961, John Wiley
and Sons, New York, New York.
6. Terlecky, P. M., M. A. Bronstein, and D. W. Goupil, Asbestos
in the Ore Mining and Dressing Industry: Identification,
Analysis, Treatment Technology, Health Aspects, and Field
Sampling Results, U.S. Environmental Protection Agency, Contract
Number 68-01-3281, 20 July 1979, page 7.
543
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SECTION XVI
GLOSSARY
absorption: The process by which a liquid is drawn into and
tends to fill permeable pores in a porous solid body; also the
increase in weight of a porous solid body resulting from the
penetration of liquid into its permeable pores.
acid copper: Copper electrode deposited from an acid solution of
a copper salt, usually copper sulfate.
acid cure: In uranium extraction, sulfation of moist ore before
leach.
acid leach: (a) Metallurgical process for dissolution of values
by means of acid solution (used on the sandstone ores of low lime
content); (b) In the copper industry, a technology employed to
recover copper from low grade ores and mine dump materials when
oxide (or mixed oxide-sulfide, or low grade sulfide)
mineralization is present, by dissolving the copper minerals with
either sulfuric acid or sulfuric acid containing ferric iron.
Four methods of leaching are employed: dump, heap, in-situ, and
vat (see appropriate definitions).
acid mine water: (a) Mine water which contains free sulfuric
acid, mainly due to the weathering of iron pyrites; (b) Where
sulfide minerals break down under the chemical influence of
oxygen and water, the mine water becomes acidic and can corrode
ironwork.
activator, activating agent: A substance which when added to a
mineral pulp promotes flotation in the presence of a collecting
agent. It may be used to increase the floatability of a mineral
in a froth, or to reflect a depressed (sunk mineral).
adit: (a) A horizontal or nearly horizontal passage driven from
the surface for the working or dewatering of a mine; (b) A
passage driven into a mine from the side of a hill.
adsorption: The adherence of dissolved, colloidal, or finely
divided solids on the surface of solids with which they are
brought into contact.
aeroflocs: Synthetic water-soluble polymers used as flocculating
agents.
all sliming: (a) Crushing all the ore in a mill to so fine a
state that only a small percentage will fail to pass through a
200-mesh screen; (b) Term used for treatment of gold ore which is
ground to a size sufficiently fine for agitation as a cyanide
pulp, as opposed to division into coarse sands for static
leaching and fine slimes for agitation.
545
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alluminothermic process: The reduction of oxides in an
exothermic reaction with finely divided aluminum.
alluvial deposit; placer deposit: Earth, sand, gravel or other
rock or mineral materials transported by and laid down by flowing
water. Alluvial deposits generally take the form of (1) surface
deposits; (2) river deposits; (3) deep leads; and (4) shore
deposits.
alunite: A basic potassium aluminum sulfate, KA13_(OH) 6(S04_) 2.
Closely resembles kaolinite and occurs in similar locations.
amalgamation: The process by which mercury is alloyed with some
other metal to produce amalgam. It was used extensively at one
time for the extraction of gold and silver from pulverized ores,
now is largely superseded by the cyanide process.
AN-FO - Ammonium nitrate: Fuel oil blasting agents.
asbestos minerals: Certain minerals which have a fibrous
structure, are heat resistant, chemically inert and possessing
high electrical insulating qualities. The two main groups are
serpentine and amphiboles. Chrysotile (fibrous serpentine, 3MgO
. 2SiO£ . 2H20) is the principal commercial variety. Other
commercial varieties are amosite, crocidolite, actinolite,
anthophyllite, and tremolite.
azurite: A blue carbonate of copper, Cu3_(C03_) 2(OH) 2_,
crystallizing in the monoclinic system. Found as an alteration
product of chalcopyrite and other sulfide ores of copper in the
upper oxidized zones of mineral veins.
bastnasite; bastnaesite: A greasy, wax-yellow to reddish-brown
weakly radioactive mineral, (Ce,La) (C03_)F, most commonly found
in contact zones, less often in pegmatites.
bauxite: (a) A rock composed of aluminum hydroxides, essentially
A12_03_ . 2H2_0. The principal ore of aluminum; also used
collectively for lateritic aluminous ores. (b) Composed of
aluminum hydroxides and impurities in the form of free silica,
clay, silt, and iron hydroxides. The primary minerals found in
such deposits are boehmite, gibbsite, and diaspore.
Bayer Process: Process in which impure aluminum in bauxite is
dissolved in a hot, strong, alkalai solution (normally NaOH) to
form sodium aluminate. Upon dilution and cooling, the solution
hydrolyzes and forms a precipitate of aluminum hydroxide.
bed: The smallest division of a stratified series and marked by
a more or less well-defined divisional plane from the materials
above and below.
546
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benef iciation: (a) The dressing or processing of ores for the
purpose of (1) regulating the size of a desired product, (2)
removing unwanted constituents, and (3) improving the quality,
purity, assay grade of a desired product/ (b) Concentration or
other preparation of ore for smelting by drying, flotation, or
magnetic separation.
Best Available Technology Economically Achievable (BAT): The
level of technology applicable to effluent limitations to be
achieved by 1 July 1983, for industrial discharges to surface
waters as defined by Section 301(b)(l)(A) of the Act.
Best Practicable Control Technology Currently Available (BPT) :
The level of technology applicable to effluent limitations to be'
achieved by 1 July 1977, for industrial discharges to surface
waters as defined by Section 301(b)(l)(A) of the Act.
byproduct: A secondary or additional product.
carbon absorption: A process utilizing the efficient absorption
characteristics of activated carbon to remove both dissolved and
suspended substances.
carnotite: A bright yellow uranium mineral, K2(U02) 2(VO^)2 .
3H2_0.
cationic collectors: In flotation, amines and related organic
compounds capable of producing positively charged hydrocarbon-
bearing ions for the purpose of floating miscellaneous minerals,
especially silicates.
cationic reagents: In flotation, surface active substances which
have the active constituent in the positive ion. Used to
flocculate and to collect minerals that are not flocculated by
the reagents, such as oleic acid or soaps, in which the surface-
active ingredient is the negative ion.
cement copper: Copper precipitated by iron from copper sulfate
solutions.
cerium metals: Any of a group of rare-earth metals separable as
a group from other metals occurring with them and in addition to
cerium includes lanthanum, praseodymium, neodymium, promethium,
samarium and sometimes europium.
cerium minerals: Rare earths; the important one is monazite.
chalcocite: Copper sulfide, Cu2S.
chalcopyrite: A sulfide of copper and iron, CuFeS2..
chert: Cryptocrystalline silica, distinguished from flint by
flat fracture, as opposed to conchoidal fracture.
547
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chromite: Chrome iron ore, FeCr^04_.
chrysocolla-. Hydrated copper silicate, CuSiOiS . 2H20.
chrysotile: A metamorphic mineral, an asbestos, the fibrous
variety of serpentine. A silicate of magnesium, with silica
tetrahedra arranged in sheets.
cinnabar: Mercury sulfide, HgS.
claim: The portion of mining ground held under the Federal and
local laws by one claimant or association, by virtue of one
location and record. A claim is sometimes called a "location."
clarification: (a) The cleaning of dirty or turbid liquids by
the removal of suspended and colloidal matter; (b) The
concentration and removal of solids from circulating water in
order to reduce the suspended solids to a minimum; (c) In the
leaching process, usually from pregnant solution, e.g., gold-rich
cyanide prior to precipitation.
classifier: (a) A machine or device for separating the
constituents of a material according to relative sizes and
densities thus facilitating concentration and treatment.
Classifiers may be hydraulic or surface-current box classifiers.
Classifiers are also used to separate sand from slime, water from
sand, and water from slime; (b) The term classifier is used in
particular where an upward current of water is used to remove
fine particles from coarser material; (c) In mineral dressing,
the classifier is a device that takes the ball-mill discharge and
separates it into two portions—the finished product which is
ground as fine as desired, and oversize material.
coagulation: The binding of individual particles to form floes
or agglomerates and thus increase their rate of settlement in
water or other liquid (see also flocculate).
coagulator: A soluble substance, such as lime, which when added
to a suspension of very fine solid particles in water causes
these particles to adhere in clusters which will settle easily.
Used to assist in reclaiming water used in flotation.
collector: A heteropolar compound containing a hydrogen-carbon
group and an ionizing group, chosen for the ability to adsorb
selectively in froth flotation processes and render the adsorbing
surface relatively hydrophobic. A promoter.
columbite; tantalite; niobite:
(columbium), tantalum, ferrous
granites and pegmatites, (Fe, Mn)
A natural
iron, and
Nb, Ta) 206
oxide of
manganese,
niobium
found in
concentrate: (a) In mining, the product of concentration; (b) To
separate ore or metal from its containing rock or earth; (c) The
548
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enriched ore after removal of waste in a beneficiation mill, the
clean product recovered in froth flotation.
concentration: Separation and accumulation of economic minerals
from gangue.
concentrator: (a) A plant where ore is separated into values
(concentrates) and rejects (tails). An appliance in such a"
plant, e.g., flotation cell, jig, electromagnet, shaking table.
Also called mill; (b) An apparatus in which, by the aid of water
or air and specific gravity, mechanical concentration of ores is
performed.
conditioners: Those substances added to the pulp to maintain the
proper pH to protect such salts as NaCN, which would decompose in
an acid circuit, etc. Na2C02[ and CaO are the most common
conditioners.
conditioning: Stage of froth-flotation process in which the
surfaces of the mineral species present in a pulp are treated
with appropriate chemicals to influence their reaction when the
pulp is aerated.
copper minerals: Those of the oxidized zone of copper deposits
(zone of oxidized enrichment) include azurite, chrysocolla,
copper metal, cuprite, and malachite. Those of the underlying
zone (that of secondary sulfide enrichment) include bornite,
chalcocite, chalcopyrite, covellite. The zone of primary
sulfides (relatively low in grade) includes the unaltered
minerals bornite and chalcopyrite.
crusher: A machine for crushing rock or other materials. Among
the various types of crushers are the ball-mill, gyratory
crusher, Hadsel mill, hammer mill, jaw crusher, rod mill, rolls,
stamp mill, and tube mill.
cuprite: A secondary copper mineral, Cu20.
cyanidation: A process of extracting gold and silver as cyanide
slimes from their ores by treatment with dilute solutions of
potassium cyanide and sodium cyanide.
cyanidation vat: A large tank, with a filter bottom, in which
sands are treated with sodium cyanide solution to dissolve out
gold.
cyclone: (a) The conical-shaped apparatus used in dust
collecting operations and fine grinding applications; (b) A
classifying (or concentrating) separator into which pulp is fed,
so as to take a circular path. Coarser and heavier fractions of
solids report at the apex of long cone while finer particles
overflow from central vortex.
549
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daughter: Decay product formed when another element undergoes
radioactive disintegration.
decant structure: Apparatus for removing clarified water from
the surface layers of tailings or settling ponds. Commonly used
structure include decant towers in which surface waters flow over
a gate (adjustable in height) and down the tower to a conduit
generally buried beneath the tailings, decant weirs over which
water flows to a channel external to the tailings pond, and
floating decant barges which pump surface water out of the pond.
dense-media separation: (a) Heavy media separation, or sink
float. Separation of heavy sinking from light floating mineral
particles in a fluid of intermediate density; (b) Separation of
relatively light (floats) and heavy ore particles (sinks), by
immersion in a bath of intermediate density.
Denver cell: A flotation cell of the subaeration type, in wide
use. Design modifications include receded-disk, conical-disk,
and multibladed impellers, low-pressure air attachments, and
special froth withdrawal arrangements.
Denver jig: Pulsion-suction diaphragm jig for fine material, in
which makeup (hydraulic) water is admitted through a rotary valve
adjustable as to portion of jigging cycle over which controlled
addition is made.
deposit: Mineral or ore deposit is used to designate a natural
occurrence of a useful mineral or an ore, in sufficient extent
and degree of concentration to invite exploitation.
depressing agent; depressor: In the froth float ion process, a
substance which reacts with the particle surface to render it
less prone to stay in the froth, thus causing it to wet down as a
tailing product (contrary to activator).
detergents, synthetic: Materials which have a cleansing action
like soap but are not derived directly from fats and oils. Used
in ore flotation.
development work: Work undertaken to open up ore bodies as
distinguished from the work of actual ore extraction or
exploratory work.
dewater: To remove water from a mine usually by pumping,
drainage or evaporation.
differential flotation: Separating a complex ore into two or
more valuable minerals and gangue by flotation; also called
selective flotation. This type of flotation is made possible by
the use of suitable depressors and activators.
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discharge: Outflow from a pump, drill hole, piping system,
channel, weir or other discernible, confined or discrete
conveyance (see also point source).
dispersing agent: Reagent added to flotation circuits to prevent
flocculation, especially of objectionable colloidal slimes.
Sodium silicate is frequently added for this purpose.
dredge/ dredging: A large floating contrivance for underwater
excavation of materials using either a chain of buckets, suction
pumps, or other devices to elevate and wash alluvial deposits and
gravel for gold, tin, platinum, heavy minerals, etc.
dressing: Originally referred to the picking, sorting, and
washing of ores preparatory to reduction. The term now includes
more elaborate processes of milling and concentration of ores.
drift mining: A term applied to working alluvial deposits by
underground methods of mining. The paystreak is reached through
an adit or a shallow shaft. Wheelbarrows or small cars may be
used for transporting the gravel to a sluice on the surface.
dump leaching: Term applied to dissolving and recovering
minerals from subore-grade materials from a mine dump. The dump
is irrigated with water, sometimes acidified, which percolates
into and through the dump, and runoff from the bottom of the dump
is collected, and a mineral in solution is recovered by chemical
reaction. Often used to extract copper from low grade, waste
material of mixed oxide and sulfide mineralization produced in
open pit mining.
effluent: The wastewater discharged from a point source to
navigable waters.
electrowinning: Recovery of a metal from an ore by means of
electrochemical processes, i.e., deposition of a metal on an
electrode by passing electric current through an electrolyte.
eluate: Solutions resulting from regeneration (elution) of ion
exchange resins.
eluent: A solution used to extract collected ions from an ion
exchange resin or solvent and return the resin to its active
state.
exploration: Location of the presence of economic deposits and
establishing their nature, shape, and grade and the investigation
may be divided into (1) preliminary, and (2) final.
extraction: (a) The process of mining and removal of ore from a
mine. (b) The separation of a metal or valuable mineral from an
ore or concentrate, (c) Used in relation to all processes that
are used in obtaining metals from their ores. Broadly, these
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processes involve the breaking down of the ore both mechanically
(crushing) and chemically (decomposition), and the separation of
the metal from the associated gangue.
ferruginous: Containing iron.
ferruginous chert: A sedimentary deposit consisting of
chalcedony or of fine-grained quartz and variable amounts of
hematite, magnetite, or limonite.
ferruginous deposit: A sedimentary rock containing enough iron
to justify exploitation as iron ore. The iron is present, in
different cases, in silicate, carbonate, or oxide form, occurring
as the minerals chamosite, thuringite, siderite, hematite,
limonite, etc.
flask: A unit of measurement for mercury; 76 pounds.
flocculant: An agent that induces or promotes flocculation or
produces floccules or other aggregate formation, especially in
clays and soils.
flocculate: To cause to aggregate or to coalesce into small
lumps or loose clusters, e.g., the calcium ion tends to
flocculate clays.
flocculating agent; flocculant: A substance which produces
flocculation.
flotation: The method of mineral separation in which a froth
created in water by a variety of reagents floats some finely
crushed minerals, whereas other minerals sink.
flotation agent: A substance or chemical which alters the
surface tension of water or which makes it froth easily. The
reagents used in the flotation process include pH regulators,
slime dispersants, resurfacing agents, wetting agents,
conditioning agents, collectors, and frothers.
friable: Easy to break, or crumbling naturally.
froth, foam: In the flotation process, a collection of bubbles
resulting from agitation, the bubbles being the agency for
raising (floating) the particles of ore to the surface of the
cell.
frother(s): Substances used in flotation processes to make air
bubbles sufficiently permanent principally by reducing surface
tension. Common frothers are pine oil, creyslic acid, and amyl
alcohol.
gangue: Undesirable minerals associated with ore.
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glory hole: A funnel-shaped excavation, the bottom of which is
connected to a raise driven from an underground haulage level or
is connected through a horizontal tunnel (drift) by which ore may
also be conveyed.
gravity separation: Treatment of mineral particles which
exploits differences between their specific gravities. Their
sizes and shapes also play a minor part in separation. Performed
by means of jigs, classifiers, hydrocyclones, dense media,
shaking tables, Humphreys spirals, sluices, vanners and briddles.
grinding: (a) Size reduction into relatively fine particles, (b)
Arbitrarily divided into dry grinding performed on mineral
containing only moisture as mined, and wet grinding, usually done
in rod, ball or pebble mills with added water.
heap leaching: A process used in the recovery of copper from
weathered ore and material from mine dumps. The liquor seeping
through the beds is led to tanks, where it is treated with scrap
iron to precipitate the copper from solution. This process can
also be applied to the sodium sulfide leaching of mercury ores.
heavy-media separation: See dense-media separation.
hematite: One of the most common ores of iron, Fe203_, which when
pure contains about 70% metallic iron and 30% oxygen. Most of
the iron produced in North America comes from the iron ranges of
the Lake Superior District, especially the Mesabi Range,
Minnesota. The hydrated variety of this ore is called limonite.
Huntington-Heberlein Process: A sink-float process employing a
galena medium and utilizing froth flotation as the means of
medium recovery.
hydraulic mining: (a) Mining by washing sand and soil away with
water which leaves the desired mineral, (b) The process by which
a bank of gold-bearing earth and rock is excavated by a jet of
water, discharged through the converging nozzle of a pipe under
great pressure. The debris is carried away with the same water
and discharged on lower levels into watercourses below.
hydrolysate; hydrolyzate: A sediment consisting partly of
chemically undecomposed, finely ground rock powder and partly of
insoluble matter derived from hydrolytic decomposition during
weathering.
hydrometallurgy: The treatment of ores, concentrates, and other
metal-bearing materials by wet processes, usually involving the
solution of some component, and its subsequent recovery from the
solution.
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ilmenite: An iron-black mineral, FeO . Ti02^ Resembles
magnetite in appearance but is readily distinguished by feeble
magnetic character.
in-situ leach: Leaching of broken ore in the subsurface as it
occurs, usually in abandoned underground mines which previously
employed block-caving mining methods.
ion(ic) exchange: The replacement of ions on the surface, or
sometimes within the lattice, of materials such as clay.
iron formation: Sedimentary, low grade, iron ore bodies
consisting mainly of chert and fine-grained quartz and ferric
oxide segregated in bands or sheets irregularly mingled (see also
taconite).
jaw crusher: A primary crusher designed to reduce large rocks or
ores to sizes capable of being handled by any of the secondary
crushers.
jig: A machine in which the feed is stratified in water by means
of a pulsating motion and from which the stratified products are
separately removed, the pulsating motion being usually obtained
by alternate upward and downward currents of the water.
jigging: (a) The separation of the heavy fractions of an ore
from the light fractions by means of a jig. (b) Up and down
motion of a mass of particles in water by means of pulsion.
laterite: Red residual soil developed in humid, tropical, and
subtropical regions of good drainage. It is leached of silica
and contains concentrations particularly of iron oxides and
hydroxides and aluminum hydroxides. It may be an ore of iron,
aluminum, manganese, or nickel.
launder: (a) A trough, channel, or gutter usually of wood, by
which water is conveyed; specifically in mining, a chute or
trough for conveying powdered ore, or for carrying water to or
from the crushing apparatus, (b) A flume.
leaching: (a) The removal in solution of the more soluble
minerals by percolating waters, (b) Extracting a soluble metallic
compound from an ore by selectively dissolving it in a suitable
solvent, such as water, sulfuric acid, hydrochloric acid, etc.
The solvent is usually recovered by precipitation of the metal or
by other methods.
leach ion-exchange flotation process: A mixed method of
extraction developed for treatment of copper ores not amenable to
direct flotation. The metal is dissolved by leaching, for
example, with sulfuric acid, in the presence of an ion exchange
resin. The resin recaptures the dissolved metal and is then
recovered in a mineralized froth by the flotation process.
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leach precipitation float: A mixed method of chemical reaction
plus flotation developed for such copper ores as chrysocolla and
the oxidized minerals. The value is dissolved by leaching with
acid, and the copper is reprecipitated on finely divided
particles of iron, which are then recovered by flotation,
yielding an impure concentrate in which metallic copper
predominates.
lead minerals: The most important industrial one is galena
(PbS), which is usually argentiferous. In the upper parts of
deposits the mineral may be altered by oxidation to cerussite
(PbCOS^) or anglesite (PbS04^. Usually galena occurs in intimate
association with sphalerite (ZnS).
leucoxene: A brown, green, or black variety of sphene or
titanite, CaTiSiO, occurring as monoclinic crystals. An earthy
alteration product consisting in most instances of rutile; used
in the production of titanium tetrachloride.
lime: Quicklime (calcium oxide) obtained by calcining limestone
or other forms of calcium carbonate. Loosely used for hydrated
lime (calcium hydroxide) and incorrectly used for pulverized or
ground calcium carbonate in agricultural lime and for calcium in
such expressions as carbonate of lime, chloride of lime, and lime
feldspar.
lime slurry: A form of calcium hydroxide in aqueous suspension
that contains considerable free water.
limonite: Hydrous ferric oxide FeO(OH) . nH2lO. An important ore
of iron, occurring in stalactitic, mammillary, or earthy forms of
a dark brown color, and as a yellowish-brown powder. The chief
constituent of bog iron ore.
liquid-liquid extraction, solvent extraction: A process in which
one or more components are removed from a liquid measure by
intimate contact with a second liquid, which is itself nearly
insoluble in the first liquid and dissolves the impurities and
not the substance that is to be purified.
lode: A tabular deposit of valuable mineral between definite
boundaries. Lode, as used by miners, is nearly synonymous with
the term vein as employed by geologists.
magnetic separation: The separation of magnetic materials from
nonmagnetic materials using a magnet. An important process in
the beneficiation of iron ores in which the magnetic mineral is
separated from nonmagnetic material, e.g., magnetite from other
minerals, roasted pyrite from sphalerite.
magnetic separator: A device used to separate magnetic from less
magnetic or nonmagnetic materials. The crushed material is
conveyed on a belt past a magnet.
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magnetite, magnetic iron ore: Natural black oxide of iron,
Fe3_04_. As black sand, magnetite occurs in placer deposits, and
also as lenticular bands. Magnetite is used widely as a
suspension solid in dense-medium washing of coal and ores.
malachite: A green, basic cupric carbonate, Cu2/OH) 2C03_,
crystallizing in the monoclinic system. It is a common ore of
copper and occurs typically in the oxidation zone of copper
deposits.
manganese minerals: Those in principal production are
pyrolusite, some psilomelane, and wad (impure mixture of
manganese and other oxides).
manganese nodules: The concretions, primarily of manganese
salts, covering extensive areas of the ocean floor. They have a
layer configuration and may prove to be an important source of
manganese.
manganese ore: A term used by the Bureau of Mines for ore
containing 35 percent or more manganese and may include
concentrate, nodules, or synthetic ore.
manganiferous iron ore: A term used by the Bureau of Mines for
ores containing 5 to 10 percent manganese.
manganiferous ore: A term used by the Bureau of Mines for any
ore of importance for its manganese content containing less than
35 percent manganese but not less than 5 percent manganese.
mercury minerals: The main source is cinnabar, HgS.
mill: (a) Reducing plant where ore is concentrated and/or metals
recovered, (b) Today the term has been broadened to cover the
whole mineral treatment plant in which crushing, wet grinding,
and further treatment of the ore is conducted. (c) In mineral
processing, one machine, or a group, used in comminution.
minable: (a) Capable of being mined, (b) Material that can be
mined under present day mining technology and economics.
mine: (a) An opening or excavation in the earth for the purpose
of excavating minerals, metal ores or other substances by
digging, (b) A word for the excavation of minerals by means of
pits, shafts, levels, tunnels, etc., as opposed to a quarry,
where the whole excavation is open. In general the existence of
a mine is determined by the mode in which the mineral is
obtained, and not by its chemical or geologic character. (c) An
excavation beneath the surface of the ground from which mineral
matter of value is extracted. Excavations for the extraction of
ore or other economic minerals not requiring work beneath the
surface are designated by a modifying word or phrase as: (1)
opencut mine - an excavation for removing minerals which is open
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to the weather; (2) steam shovel mine - an opencut mine in which
steam shovels or other power shovels are used for loading cars;
(3) strip mine - a stripping, an openpit mine in which the
overburden is removed from the exploited material before the
material is taken out; (4) placer mine - a deposit of sand,
gravel or talus from which some valuable mineral is extracted;
and (4) hydraulic mine - a placer mine worked by means of a
stream of water directed against a bank of sand, gravel, or
talus. Mines are commonly known by the mineral or metal
extracted, e.g., bauxite mines, copper mines, silver mines, etc.
(d) Loosely, the word mine is used to mean any place from which
minerals are extracted, or ground which it is hoped may be
mineral bearing, (e) The Federal and State courts have held that
the word mine, in statutes reserving mineral lands, included only
those containing valuable mineral deposits. Discovery of a mine:
In statutes relating to mines the word discovery is used: (1) In
the sense of uncovering or disclosing to view ore or mineral; (2)
of finding out or bringing to the knowledge the existence of ore,
or mineral, or other useful products which were unknown; and (3)
of exploration, that is, the more exact blocking out or
ascertainment of a deposit that has already been discovered. In
this sense it is practically synonymous with development, and has
been so used in the U.S. Revenue Act of 19 February 1919 (Sec.
214, subdiv. A10, and Sec. 234, subdiv. A9) in allowing
depletion of mines, oil and gas wells. Article 219 of Income and
War Excess Profits Tax Regulations No. 45, construes discovery of
a mine as: (1) The bona fide discovery of a commercially
valuable deposit of ore or mineral, of a value materially in
excess of the cost of discovery in natural exposure or by
drilling or other exploration conducted above or below the
ground; and (2) the development and proving of a mineral or ore
deposit which has been apparently worked out to be a mineable
deposit or ore, or mineral having a value in excess of the cost
of improving or development.
mine drainage: (a) Mine drainage usually implies gravity flow of
water to a point remote from mining operation, (b) The process of
removing surplus ground or surface water by artificial means.
mineral: An inorganic substance occurring in nature, though not
necessarily of inorganic origin, which has (1) a definite
chemical composition, or commonly a characteristic range of
chemical composition, and (2) distinctive physical properties, or
molecular structure. With few exceptions, such as opal
(amorphous) and mercury (liquid), minerals are crystalline
solids.
mineral processing; ore dressing; mineral dressing: The dry and
wet crushing and grinding of ore or other mineral-bearing
products for the purpose of raising concentrate grade; removal of
waste and unwanted or deleterious substances from an otherwise
useful product; separation into distinct species of mixed
minerals; chemical attack and dissolution of selected values.
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modifier(s): (a) In froth flotation, reagents used to control
alkalinity and to eliminate harmful effects of colloidal material
and soluble salts. (b) Chemicals which increase the specific
attraction between collector agents and particle surfaces, or
conversely which increase the wettability of those surfaces.
molybdenite: The most common ore of molybdenum, MoSz_.
molybdenite concentrate: Commercial molybdenite ore after the
first processing operations. Contains about 90% MoS,2 along with
quartz, feldspar, water, and processing oil.
monazite: A phosphate of the cerium metals and the principal ore
of the rare earths and thorium. Monoclinic. One of the chief
sources of thorium used in the manufacture of gas mantles. It is
a moderately to strongly radioactive mineral, (Ce, La, Y, Th)
PCM. It occurs widely disseminated as an accessory mineral in
granitic igneous rocks and gneissic metamorphic rocks. Detrital
sands in regions of such rocks may contain commercial quantities
of monazite. Thorium-free monazite is rare.
New Source Performance Standard (NSPS): Performance standards
for the industry and applicable new sources as defined by Section
306 of the Act.
niccolite: A copper-red arsenide of nickel which usually
contains a little iron, cobalt, and sulfur. It is one of the
chief ores of metallic nickel.
nickel minerals: The nickel-iron sulfide, pentlandite (Fe, Ni)
is the principal present economic source of nickel, and
garnierite (nickelmagnesium hydrosilicate) is next in economic
importance.
oleic acid: A mono-saturated fatty acid, CH3_(CH2_)_CH:CH(CH2)7_
COOH. A common component of almost all naturally occurring fats
as well as tall oil. Most commercial oleic acid is derived from
animal tallow or natural vegetable oils.
open-pit mining, open cut mining: A form of operation designed
to extract minerals that lie near the surface. Waste, or
overburden, is first removed, and the mineral is broken and
loaded. Important chiefly in the mining of ores of iron and
copper.
ore: (a) A natural mineral compound of the elements of which one
at least is a metal. Applied more loosely to all metalliferous
rock, though it contains the metal in a free state, and
occasionally to the compounds of nonmetallic substances, such as
sulfur. (b) A mineral of sufficient value as to quality and
quantity which may be mined with profit.
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ore dressing: The cleaning of ore by essentially physical means
and the removal of certain valueless portion. Synonym for
concentration. The same as mineral dressing.
ore reserve: The term usually restricted to ore of which the
grade and tonnage have been established with reasonable assurance
by drilling and other means.
oxidized ores: The alteration of metalliferous minerals by
weathering and the action of surface waters, and the conversion
of the minerals into oxides, carbonates, or sulfates.
oxidized zone: That portion of an ore body near the surface,
which has been leached by percolating water carrying oxygen,
carbon dioxide or other gases.
pegmatite: An igneous rock of coarse grain size usually found as
a crosscutting structure in a larger igneous mass of finer grain
size.
pelletizing: A method in which finely divided material is rolled
in a drum or on an inclined disk, so that the particles cling
together and roll up into small, spherical pellets.
pH modifiers: Proper functioning of a cationic or anionic
flotation reagent is dependent on the close control of pH.
Modifying agents used are soda ash, sodium hydroxide, sodium
silicate, sodium phosphates, lime, sulfuric acid, and
hydrofluoric acid.
placer mine: (a) A deposit of sand, gravel, or talus from which
some valuable mineral is extracted, (b) To mine gold, platinum,
tin or other valuable minerals by washing the sand, gravel, etc.
placer mining: The extraction of heavy mineral from a placer
deposit by concentration in running water. It includes ground
sluicing, panning, shoveling gravel into a sluice, scraping by
power scraper, excavation by dragline or extraction by means of
various types of dredging activities.
platinum minerals: Platinum, ruthenium rhodium, palladium,
osmium, and iridium are members of a group characterized by high
specific gravity, unusual resistance to oxidizing and acidic
attac.k, and high melting point.
point source: Any discernible, confined and discrete conveyance,
including but not limited to any pipe, ditch, channel, tunnel,
conduit, well, discrete fissure, container, rolling stock,
concentrated animal feeding operation, or vessel or other
floating craft, from which pollutants are or may be discharged.
pregnant solution: A value bearing solution in a
hydrometallurgical operation.
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pregnant solvent: In solvent extraction, the value-bearing
solvent produced in the solvent extraction circuit.
promoter: A reagent used in froth-flotation process, usually
called the collector.
rare-earth deposits: Sources of cerium, terbium, yttrium, and
related elements of the rare-earth's group, as well as thorium.
raw mine drainage: Untreated or unprocessed water drained,
pumped or siphoned from a mine.
reagent: A chemical or solution used to produce a desired
chemical reaction; a substance used in assaying or in flotation.
reclamation: The procedures by which a disturbed area can be
reworked to make it productive, useful, or aesthetically
pleasing.
recovery: A general term to designate the valuable constituents
of an ore which are obtained by metallurgical treatment.
reduction plant: A mill or a treatment place for the extraction
of values from ore.
roast: To heat to a point somewhat short of fuzing in order to
expel volatile matter or effect oxidation.
rougher cell: Flotation cells in which the bulk of the gangue is
removed from the ore.
roughing: Upgrading of run-of-mill feed either to produce a low
grade preliminary concentrate or to reject valueless tailings at
an early stage. Performed by gravity on roughing tables, or in
flotation in a rougher circuit.
rutile: Titanium dioxide, Ti02.
scintillation counter: An instrument used for the location of
radioactive ore such as uranium. It uses a transparent crystal
which gives off a flash of light when struck by a gamma ray, and
a photomultiplier tube which produces an electrical impulse when
the light from the crystal strikes it.
selective flotation: See differential flotation.
settling pond: A pond, natural or artificial, for recovering
solids from an effluent.
siderite: An iron carbonate, FeC03_.
slime, slimes: A material of extremely fine particle size
encountered in ore treatment.
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sludge: The precipitant or settled material from a wastewater.
slurry: (a) Any finely divided solid which has settled out as
from thickeners, (b) A thin watery suspension.
solvent extraction: See liquid-liquid extraciton.
sphalerite: Zinc sulfide, ZnS.
stibnite: An antimony sulfide, Sb2_S3_. The most important ore of
antimony.
suction dredge: (a) Essentially a centrifugal pump mounted on a
barge. (b) A dredge in which the material is lifted by pumping
through a suction pipe.
sulfide zone: That part of a lode or vein not yet oxidized by
the air or surface water and containing sulfide minerals.
surface active agent: One which modifies physical, electrical,
or.chemical characteristics of the surface of solids and also
surface tensions of solids or liquid. Used in froth flotation
(see also depressing agent, flotation agent).
tabling: Separation of two materials of different densities by
passing a dilute suspension over a slightly inclined table having
a reciprocal horizontal motion or shake with a slow forward
mot-ion and a fast return.
taconite: (a) The cherty or jaspery rock that encloses the
Mesabi iron ores in Minnesota. In a somewhat more general sense,
it designates any bedded ferruginous chert of the Lake Superior
District, (b) In Minnesota practice, is any grade of extremely
hard, lean iron ore that has its iron either in banded or well-
desseminated form and which may be hematite or magnetite, or a
combination of the two within the same ore body (Bureau of
Mines).
taconite ore: A type of highly abrasive iron ore now extensively
mined in the United States.
tailing pond: Area closed at lower end by constraining wall or
dam to ^hich mill effluents are run.
tailings: (a) The parts, or a part, of any incoherent or fluid
material separated as refuse, or separately treated as inferior
in quality or value; leavings; remainders; dregs. (b) The gangue
and other refuse material resulting from the washing,
concentration, or treatment of ground ore. (c) Those portions of
washed ore that are regarded as too poor to be treated further;
used especially of the debris from stamp mills or other ore
dressing machinery, as distinguished from concentrates.
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tall oil: The oily mixture of rosin acids, and other materials
obtained by acid treatment of the alkaline liquors from the
digesting (pulping) of pine wood. Used in drying oils, in
cutting oils, emulsifiers, and in flotation agents.
tantalite: A tantalate of iron and manganese (Fe, Mn) Ta^O,
crystallizing in the orthorhombic system.
tetrahedrite: A mineral, the part with Sb greater than As of the
tetrahedrite-tenantite series, Cu3_(Sb, As)S3^. Silver, zinc, iron
and mercury may replace part of the copper. An important ore of
copper and silver.
thickener: A vessel or apparatus for reducing the amount of
water in a pulp.
thickening: (a) The process of concentrating a relatively dilute
slime pulp into a thick pulp, that is, one containing a smaller
percentage of moisture, by rejecting liquid that is essentially
solid free. (b) The concentration of the solids in a suspension
with a view to recovering one fraction with a higher
concentration of solids than in the original suspension.
tin minerals: Virtually all th,e industrial supply comes from
cassiterite (Sn02_), though some has been obtained from the
sulfide minerals stannite, cylindrite, and frankeite. The bulk
of cassiterite comes from alluvial workings.
titanium minerals: The main commercial minerals are rutile
(Ti02.) and ilmenite (FeTi03_) .
tyuyamunite: A yellow uranium mineral (Ca(UO£)2_V04.)l . 3H20. It
is the calcium analogue of carnotite.
uraninite: Essentially U02_. It is a complex uranium mineral
containing also rare earths, radium, lead, helium, nitrogen and
other elements.
uranium minerals: More than 150 uranium bearing minerals are
known to exist, but only a few are common. The five primary
uranium-ore minerals are pitchblende, uraninite, davidite,
coffinite, and brannerite. These were formed by deep-seated hot
solutions and are most commonly found in veins or pegmatites.
The secondary uranium-ore minerals, altered from the primary
minerals by weathering or other natural processes, are carnotite,
tyuyamunite and metatyuyamunite (both are very similar to
carnotite), torbernite and metatorbernite, autunite and meta-
autunite, and uranophane.
vanadium minerals: Those most exploited for industrial use are
patronite (VS4_) , roscoelite (vanadium mica), vanadinite
(Pb_Cl (V04_) 3_), carnotite and chlorovanadinite.
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vat leach: Employs the dissolution of copper oxide minerals by
sulfuric acid from crushed, non-porous ore material placed in
confined tanks. The leach cycle is rapid and measured in days.
weir: An obstruction placed across a stream for the purpose of
diverting the water so as to make it flow through a desired
channel, which may be an opening or notch in the weir itself.
wetting agent: A substance that lowers the surface tension of
water and thus enables it to mix more readily. Also called
surface active agent.
Wilfley table: Widely used for of shaking table. A plane
rectangle is mounted horizontally and can be sloped about its
long axis. It is covered with linoleum (occasionally rubber) and
has longitudinal riffles dying at the discharge end to a smooth
cleaning area, triangular in the upper corner. Gentle and rapid
throwing motion is used on the table longitudinally. Sands,
usually classified for size range are fed continuously and worked
along the table with the aid of feedwater, and across riffles
downslope by gravity tilt adjustment, and added washwater. At
the discharge end, the sands have separated into bands, the
heaviest and smallest uppermost, the lightest and largest lowest.
xanthate: Common specific promoter used in flotation of sulfide
ores. A salt or ester of xanthic acid which is made of an
alcohol, carbon disulfite and an alkalai.
xenotime: A yttrium phosphate, YPO£, often containing small
quantities of cerium, terbium, and thorium, closely resembling
zircon in crystal form and general appearance.
yellow cake: (a) A term applied to certain uranium concentrates
produced by mills. It is the final precipitate formed in the
milling process. It is usually considered to be ammonium
diuranate, (NH4_) 2\J207_, or sodium diuranate, Na2U207_, but the
composition is variable and depends upon the precipitating
conditions, (b) A common form of triuranium octoxide, U308_, is
yellow cake, which is the powder obtained by evaporating an
ammonia solution of the oxide.
zinc minerals: The main source of zinc is sphalerite (ZnS), but
some smithsonite, hemimorphite, zincite, willemite, and
franklinite are mined.
zircon: A mineral, ZrSiO£. The chief ore of zirconium.
zircon, rutile, ilmenite, monazite: A group of heavy minerals
which are usually considered together because of their occurence
as black sand in natural beach and dune concentration.
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APPENDIX A
INDUSTRY PROCESSES
(Refer to Section III)
The following ore types are included: iron, copper, lead-zinc,
gold, silver, molybdenum, tungsten, vanadium, mercury, uranium,
antimony and titanium.
Iron Ore Milling Processes
Beneficiation of iron ore includes such operations as crushing,
screening, blending, grinding, concentrating, classifying,
briquetting, sintering and agglomerating. Beneficiation is often
done at or near the mine site. Methods selected are based on
physical and chemical properties of the crude ore. General
techniques utilized in the beneficiation of iron ore are
illustrated in Figure A-l. Processes enhance either the chemical
or physical characteristics of the crude ore to make more
desirable feed for the blast furnace. Beneficiation methods have
been developed to upgrade 20 to 30 percent iron 'taconite' ores
into high-grade materials.
Physical concentrating processes, such as washing, remove
unwanted sand, clay, or rock from crushed or screened ore. For
those ores not amenable to simple washing operations, other phys-
ical methods are used such as jigging, heavy-media separation,
flotation, and magnetic separation. Jigging involves stratifica-
tion of ore and gangue by utilizing pulsating water currents.
Heavy-media separation employs water suspension of ferrosilicon
whereby iron ore particles sink while the majority of gangue
(quartz, etc.) floats. The flotation process uses air bubbles
attached to iron particles conditioned with flotation reagents to
separate iron from the gangue. Magnetic separation techniques
are used on ores containing magnetite.
At the present time, there are only three iron ore flotation
plants in the United States. Figure A-2 illustrates a typical
flowsheet used in an iron ore flotation circuit, while Table A-l
lists types and amounts of flotation reagents used per ton of ore
processed. Various flotation methods which utilize these rea-
gents are listed in Table A-2. The most commonly adopted flow-
sheet for the beneficiation of low grade magnetic taconite ores
is illustrated in Figure A-3. Low grade ores containing magne-
tite are very susceptible to concentrating processes, yielding a
high quality blast furnace feed. Higher grade iron ores con-
taining hematite cannot be upgraded much above 55 percent iron.
Figure A-4 illustrates the beneficiation of a fine-grained
hematite ore.
Agglomeration, which follows concentration processes, increases
the particle size of iron ore and reduces "fines" which normally
would be lost in the flue gases. Agglomerating methods include
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sintering, palletizing, briquetting, and nodulizing. Sintering
involves the mixing of small portions of coke and limestone with
the iron ore, followed by combustion. A granular, coarse, porous
product is formed. Pelletizing involves the formation of pellets
or balls composed of iron ore fines, followed by heating (Figure
A-5 illustrates a typical pelletizing operation). Hot ore
briquetting requires no binder, is less sensitive to changes in
feed composition, requires little or no grinding and requires
less fuel than sintering. Small or large lumps of regular shape
are formed. Nodules or lumps (nodulizing) are formed when ores
are charged into a rotary kiln and heated to incipient fusion
temperatures.
Copper Ore Milling Processes
Processing of copper ores may involve hydrometallurgical or
physical-chemical separation from the gangue material. A general
scheme of methods employed for recovery of copper from ores is
shown in Figure A-6. These methods include dump, heap, vat and
in-situ leaching, and froth flotation.
Cement copper is produced from dump, heap and in-situ leaching
and cathode copper is produced by electrowinning the pregnant
solution from a vat leach. Major copper areas employing dump,
heap and in-situ leaching are shown in Figure A-7.
Copper bearing froth from the froth flotation process is
thickened, filtered and sent to a smelter whereby blister copper
(98 percent Cu) is produced. The blister copper is then sent to
a refinery which produces pure copper (99.88 to 99.9 percent Cu)
for market.
One combination of the hydrometallurgical and physical-chemical
processes, termed LPF (leach-precipitation-flotation) has enabled
the copper industry to process oxide and sulfide minerals
efficiently. Also, tailings from the vat leaching process, if
they contain significant sulfide copper, can be sent to the
flotation circuit to float copper sulfide, while the vat leach
solution undergoes iron precipitation or electrowinning to
recover copper dissolved from oxide ores by acid.
Lead-Zinc Ore Milling Processes
Generally, lead-zinc ores are not of high enough grade to be
smelted directly, therefore it is sent through the milling pro-
cess first. In most cases, the only process utilized is froth
flotation, but in some cases, preliminary gravity separation is
practiced prior to flotation. The general milling procedure is
to crush the ore and then grind it, in a closed circuit with rod
mills, ball mills and classifying equipment, to a small enough
size to allow the ore minerals to be freed from the gangue.
Chemical reagents are then added which, in the presence of forced
air bubbles, produce selective flotation and separation of the
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desired ore minerals. In some cases, the reagents used in the
flotation process are added in the mill; in other cases, the fine
material from the mill flows to a conditioner (mixing tank),
where the reagents are added. The particular reagents utilized
are a function of the mineral concentrates to be recovered. The
specific choice of reagents used at a facility is usually the
result of determining empirically which reagents yield the opti-
mum mineral values versus reagent costs. In general, lead and
zinc as well as copper sulfide flotations are run at elevated pH
(8.5 to 11, generally) levels so that frequent pH adjustments
with hydrated lime (CaOHi2) are common. Other reagents commonly
used are:
Reagent Purpose
Methyl Isobutyl-carbinol Frother
Propylene Glycol Methyl Ether Frother
Long-Chain Aliphatic Alcohols Frother
Pine Oil Frother
Potassium Amyl Xanthate Collector
Sodium Isopropol Xanthate Collector
Sodium Ethyl Xanthate Collector
Dixanthogen Collector
Isopropyl Ethyl Thionocarbonate Collectors
Sodium Diethyl-dithiophosphate Collectors
Zinc Sulfate Zinc Depressant
Sodium Cyanide Zinc Depressant
Copper Sulfate Zinc Activant
Sodium Dichromate Lead Depressant
Sulfur Dioxide Lead Depressant
Starch Lead Depressant
Lime pH Adjustment
The finely ground ore slurry is introduced into a series of
flotation cells, where the slurry is agitated and air is
introduced. The desired minerals are rendered hydrophobia (non-
water-accepting) by surface coating with appropriate reagents.
Usually, several cells are operated in a countercurrent flow
pattern, with the final concentrate being floated off the last
cell (cleaner) and the tails being removed from the first or
rougher cells.
In many cases, more than one mineral is recovered. In such
cases, differential flotation is practiced. The flow diagram in
Figure A-8 depicts a typical differential flotation process for
recovery of lead and zinc sulfides. Chemicals which induce
hydrophilic (affinity-for-water) behavior by surface interaction
are added to prevent one of the minerals from floating in the
initial separation. The underflow of tailings from this separa-
tion is then treated with a chemical which overcomes the depres-
sing effect and allows the flotation of the other mineral.
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The floated concentrates are dewatered (usually by thickening and
filtration), and the final concentrate—which contains some
residual water—is eventually shipped to a smelter for metal
recovery. The liquid overflow from the concentrate thickeners is
typically recycled in the mill.
After the recovery of the desirable minerals, a large volume of
tailings or gangue material remains as underflow from the last
rougher cell in the flow scheme. These tails are typically
adjusted to a slurry suitable for hydraulic transport to -the
treatment facility, i.e., tailing pond. In some cases, the
coarse tailings are removed by a cyclone separator and then
pumped in to the mine for backfilling.
The tailings from a lead/zinc flotation mill contain residual
solids from the original ore which has been finely ground to
allow mineral recovery. The tailings also contain dissolved
solids and excess mill reagents. In cases where the mineral
content of the ore varies, excess reagents will undoubtedly be
present when the ore grade drops suddenly, conversely lead and
zinc will escape with the tails if high-grade ore creates a
reagent-starved system. Accidental spilling of the chemical
reagents used are another source of adverse discharges from a
mill.
Figure A-8 depicts a typical lead-zinc ore mining and processing
operation.
Gold Ore Milling Processes
Milling practices applicable to the processing and recovery of
gold and gold-containing ores are cyanidation, amalgamation,
flotation, and gravity concentration. All of these processes
have been used in the beneficiation of ore mined from lode
deposits. Placer operations, however, employ only gravity
methods which in the past were sometimes used in conjunction with
amalgamation.
Prior to 1970, the amalgamation process was used to recover
nearly 1/4 of the gold produced domestically. Since that time,
environmental concerns have caused restricted use of mercury. As
a result, the percent of gold produced which was recovered by the
amalgamation process dropped from 20.3 percent in 1970 to 0.3
percent in 1972. At the same time, the use of cyanidation
processes was increasing. In 1970, 36.7 percent of the gold
produced domestically was recovered by cyanidation, and this
increased to 54.6 percent in 1972.
The amalgamation process as currently practiced (used by a single
mill in Colorado) involves crushing and grinding of the lode ore,
gravity separation of the gold-bearing black sands by jigging,
and final concentration of the gold by batch amalgamation of the
sands in a barrel amalgamator. In the past, amalgamation of lode
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ore has been performed in either the grinding mill, on plates, or
in special amalgamators. Placer gold/silver-bearing gravels are
beneficiated by gravity methods, and, in the past, the precious
metal-bearing sands generally were batch amalgamated in barrel
amalgamators. However, amalgamation in specially designed sluice
boxes were also practiced.
There are basically four methods of cyanidation currently being
used in the United States: heap leaching, vat leaching,
agitation leaching, and the recently developed carbon-in-pulp
process. Heap leaching is a process used primarily for the
recovery of gold from low-grade ores. This is an inexpensive
process and, as a result, has also been used recently to recover
gold from old mine waste dumps. Higher grade ores are often
crushed, ground, and vat leached or agitated/leached to recover
the gold.
In vat leaching, a vat is filled with the ground ore (sands)
slurry, water is allowed to drain off, and the sands are leached
from the top with cyanide, which solubilizes the gold (Figure A-
9). Pregnant cyanide solution is collected from the bottom of
the vat and sent to a holding tank. In agitation leaching, the
cyanide solution is added to a ground ore pulp in thickeners, and
the mixture is agitated until solution of the gold is achieved
(Figure A-10). The cyanide solution is collected by decanting
from the thickeners.
Cyanidation of slimes, generated during wet grinding, is cur-
rently being done by a recently developed process, carbon-in-pulp
(Figure A-9). The slimes are mixed with a cyanide solution in
large tanks, and the solubilized gold cyanide is collected by
adsorption onto activated charcoal. Gold is stripped from the
charcoal using a small volume of hot caustic; an electrowinning
process is used for final recovery of the gold in the mill.
Bullion is subsequently produced at a refinery.
Gold in the pregnant cyanide solutions from heap, vat, or agitate
leaching processes is recovered by precipitation with zinc dust.
The precipitate is collected in a filter press and sent to a
smelter for the production of bullion.
Recovery of gold by flotation processes is limited, and less than
3 percent of the gold produced in 1972 was recovered in this
manner. This method employs a froth flotation process to float
and collect the gold-containing minerals (Figure A-l1). The one
operation that uses this method, further processes tailings from
the flotation circuit by the agitation/cyanidation method to
recover the residual gold values.
Gold has historically been recovered from placer gravels by
purely physical means. Present practice involves gravity separa-
tion, which is normally accomplished in a sluice box. Typically,
a sluice box consists of an open box in which a simple rectangu-
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lar sluice plate is mounted on a downward incline. To effect the
separation of gold from gravel or sand, a perforated metal sheet
is fitted on the bottom of the loading box and riffle structures
are mounted on the bottom of the sluice plate. These riffles may
consist of wooden strips or steel or plastic plates which are
angled away from the direction of flow in a manner designed to
create pockets and eddy currents for the collection and retention
of gold.
During actual sluicing operations, pay gravels (i.e., goldbearing
gravels) are loaded into the upper end of the sluice box and
washed down the sluice plate with water, which enters at right
angles to (or against the direction of) gravel feed. Density
differences allow the particles of gold to settle and become
entrapped in the spaces between the riffle structures, while the
less-dense gravel and sands are washed down the sluice plate.
Eddy currents keep the spaces between riffle structures free of
sand and gravel but are not strong enough to wash out the gold
which collects there.
Other types of equipment which may be employed in physical
separation operations include jigs, tables, and screens.
However, this equipment is typically found only at dredging
operations.
Cleanup of gold recovered by gravity methods is normally
accomplished with small (102 cm (40 in.)) sluices, screens, and
finally, by hand-picking impurities from the gold.
Silver Ore Milling Processes
Present extractive metallurgy for silver was developed over a
period of more than 100 years. Initially, silver, as the major
product, was recovered from rich oxidized ores by relatively
crude methods. As the ores became leaner and more complex, an
improved extractive technology was developed. Today, silver pro-
duction is predominantly as a byproduct, and is largely related
to the production of lead, zinc, and copper from the processing
of sulfide ores by froth flotation and smelting. Free-milling,
easily liberated gold/silver ores, processed by amalgamation and
cyanidation, now contribute only 1 percent of the domestic silver
produced. Primary sulfide ores, processed by flotation and
smelting, account for 99 percent.
Selective froth flotation processing can effectively and effi-
ciently beneficiate almost any type and grade of sulfide ore.
This process employs various well-developed reagent combinations
and conditions to enable the selective recovery of many different
sulfide minerals in separate concentrates of high quality. The
reagents commonly used in the process are generally classified as
collectors, promoters, modifiers, depressants, activators, and
frothing agents. Essentially, these reagents are used in combin-
ation to cause the desired sulfide mineral to float and be
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collected in a froth while the undesired minerals and gangue
sink. Practically all the ores presently milled require fine
grinding to liberate the sulfide minerals from one another and
from the gangue minerals.
A circuit which exemplifies the current practice of froth
flotation for the primary recovery of silver from silver ores or
complex ores is shown in Figure A-12. Primary recovery of silver
occurs mainly from the mineral tetrahedrite, (Cu, Fe, In, Ag)
ilsbisil- A tetrahedrite concentrate contains approximately 25
to 32 percent copper in addition to the 25.72 to 44.58 kilograms
per metric ton (750 to 1,300 troy ounces per ton) of silver. A
low-grade (3.43 kg per metric ton; 100 troy ounces per ton)
silver/pyrite concentrate is produced at one mill. Antimony may
comprise up to 18 percent of the tetrahedrite concentrate and may
or may not be extracted prior to shipment to a smelter.
Various other silver-containing minerals are recovered as
byproducts of primary copper, lead, and/or zinc operations.
Where this occurs, the usual practice is to ultimately recover
the silver from the base-metal flotation concentrates at the
smelter or refinery.
Molybdenum Ore Milling Processes
The only commercially important ore of molybdenum is molybdenite,
MoS2_. It is universally concentrated by flotation. Significant
quantities of molybdenite concentrate are recovered as a
byproduct in the milling of copper and tungsten ores.
Flotation concentration has become a mainstay of the ore milling
industry. Because it is adaptable to very fine particle sizes
(less than 0.01 mm, or 0.0004 inch), it allows high rates of
recovery from slimes which are inevitably generated in crushing
and grinding and are not generally amenable to physical
processing. As a physicochemical surface phenomenon, it can
often be made highly specific, allowing production of high-grade
concentrates from very-low-grade ore. Its specificity also
allows separation of different ore minerals (e.g., CuS and MoS2_)
where desired, and operation with minimum reagent consumption
since reagent interaction is typically only with the particular
materials to be floated or depressed.
The major operating plants in the industry recover molybdenite by
flotation. Vapor oil is used as the collector, and pine oil is
used as a frother. Lime is used to control pH of the mill feed
and to maintain an alkaline circuit. In addition, Nokes reagent
and sodium cyanide are used to prevent flotation of galena and
pyrite with the molybdenite. A generalized, simplified flowsheet
for an operation recovering only molybdenite is shown in Figure
A-13. Water use in this operation currently amounts to
approximately 1.8 tons of water per ton of ore processed,
essentially all of which is process water. Reclaimed water from
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thickeners at the mill site (shown on the flowsheet) amounts to
only 10 percent of total use.
Where byproducts are recovered with molybdenite, a somewhat more
complex mill flowsheet results, although the molybdenite recovery
circuits remain quite similar. A very simplified flow diagram
for such an operation is shown in Figure A-14. Pyrite flotation
and monazite flotation are accomplished at acid pH (4.5 and 1.5,
respectively), thereby increasing the likelihood of solubilizing
heavy metals. Flow volumes at those locations in the circuit are
low, however, and neutralization occurs upon combination with the
main mill water flows for delivery to the tailing ponds. Water
flow for this operation amounts to approximately 2.3 tons per ton
of ore processed, nearly all of which is process water in contact
with the ore. Essentially 100 percent recycle of mill water from
the tailing ponds at this mill is prompted by limited water
availability as well as by environmental considerations.
Tungsten Ore Milling Processes
Commercially important tungsten ores include the scheelite
(CaW04_) and wolframite series, wolframite ((Fe, MN) W04_),
ferberite (FeW04_), and huebnerite (MnW04_). Concentration is by a
wide variety of techniques. Gravity concentration, by jigging,
tabling, or sink/float methods, is frequently employed. Because
sliming due to the high friability of scheelite ore (most U.S.
ore is scheelite) reduces recovery by gravity techniques, fatty-
acid flotation may be used to increase recovery. Leaching may
also be employed as a major beneficiation step and is frequently
practiced to lower the phosphorus content of concentrates. Ore
generally contains about 0.6 percent tungsten, and concentrates
containing about 70 percent W03_ are produced. A tungsten
concentrate is also produced as a byproduct of molybdenum milling
at one operation in a process involving gravity separation,
flotation, and magnetic separation.
Figure A-15 depicts a simplified flow diagram for a small
tungsten concentrator.
Vanadium Ore Processes
Eighty-six percent of vanadium oxide production has recently been
used in the preparation of ferrovanadium. Although a fair share
of U.S. vanadium production is derived as a byproduct of the
mining of uranium, there are other sources of vanadium ores. The
environmental considerations at mine/mill operations not
involving radioactive constituents are fundamentally different
from environmental considerations important to uranium
operations, and it seems appropriate to consider the former
operation separately. Vanadium is considered as part of this
industry segment: (a) because of the similarity of
nonradioactive vanadium recovery operations to the processes used
for other ferroalloy metals and (b) because, in particular,
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hydrometallurgical processes like those used in vanadium recovery
are becoming more popular in SIC 1061.
Vanadium is chemically similar to columbium (niobium) and
tantalum, and ores of these metals may be beneficiated in the
same type of process used for vanadium. There is also some
similarity to tungsten, molybdenum, and chromium.
Recovery of vanadium phosphate rocks in Idaho, Montana, Wyoming,
and Utah—which contain about 28 percent P205_, 0.25 percent V2_05_,
and some Cr, Ni, and Mo—yields vanadium as a byproduct of
phosphate fertilizer production. Ferrophosphate is first
prepared by smelting a charge of phosphate rock, silica, coke,
and iron ore (if not enough iron is present in the ore). The
product when separated from the slag typically contains 60 per-
cent iron, 25 percent phosphorus, 3 to 5 percent chromium, and 1
percent nickel. It is pulverized, mixed with soda ash (Na2C03_)
and salt, and toasted at 750 to 800 degrees Celsius (1382 to 1472
degrees Fahrenheit). Phosphorus, vanadium, and chromium are
converted to water-soluble trisodium phosphate, sodium
metavanadate, and sodium chromate, while the iron remains in
insoluble form and is not extracted in a water leach following
the roast.
Phosphate values are removed from the leach in three stages of
crystallization. Vanadium can be recovered as V2_05_ (redcake) by
acidification, and chromium is precipitated as lead chromate. By
this process, 85 percent of the vanadium, 65 percent of the
chromium, and 91 percent of the phosphorus can be extracted.
Another, basically non-radioactive, vanadium ore, with a grade of
1 percent V2_05_, is found in a vanidiferous, mixed-layer
montmorillonite/illite and geothite/montroseite matrix. This ore
is recovered by salt roasting, following extrusion of pellets, to
yield sodium metavanadate, which is concentrated by solvent
extraction. Slightly soluble ammonium vanadate is precipitated
from the stripping solution and calcined to yield vanadium
pentoxide. A flow chart for this process is shown in Figure A-
16.
Mercury Ore Mill ing Processes
The principal mineral source of mercury is cinnabar (HgS). The
domestic industry has been centered in California, Nevada, and
Oregon. Mercury has also been recovered from ore in Arizona,
Alaska, Idaho, Texas, and Washington and is recovered as a
byproduct from gold ore in Nevada and zinc ore in New York.
Until recently, the typical practice of the industry has been to
feed mercury ore directly into rotary kilns for recovery of
mercury by roasting. This has been such an efficient method that
extensive beneficiation is precluded. However, with the deple-
tion of high grade ores, concentration of low-grade mercury ores
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is becoming more important. The ore may be crushed and sometimes
screened to provide a feed suitable for furnacing. Gravity con-
centration is also done in a few cases, but its use is limited
since mercury minerals crush more easily and more finely than
gangue rock.
Flotation is the most efficient method for beneficiating mercury
ores when beneficiation is practiced. An advantage of flotation,
especially for low-grade material, is the high ratio of
concentration that results. This permits proportionate
reductions in the size and costs of the final mercury extraction
process. Only recently has flotation of mercury been practiced
in the United States. During 1975, a single mill, located in
Nevada, began operation to beneficiate mercury ore by this method
(Figure A-17). The concentrate produced is furnaced at the same
facility to recover elemental mercury. The ore, which averages
4.8 kg of mercury per metric ton (9.5 Ib/short ton), is obtained
from a nearby open-pit mine/ the major ore minerals present are
cinnabar (HgS) and corderoite (Hg3S2Cl2_).
Uranium Ore Milling Processes
Blending, Crushing and Roasting. Ore from the mine can be quite
variable in consistency and grade. Procedures have been
developed to weigh and radiometrically assay the ores. This is
done to achieve uniform grade and consistency.
Ore high in vanadium is sometimes roasted with sodium chloride
after crushing. This converts insoluble heavy-metal vanadates
(vanadium complex) and carnotite to more soluble sodium vanadate,
which is then extracted with water. Ores high in organics may be
roasted to carbonize and oxidize the organics and prevent clog-
ging of hydrometallurgical processes. Clay bearing ores attain
improved filtering and settling characteristics by roasting at
300 degrees Celsius (572 degrees Fahrenheit).
Grinding. Ore is ground less than 0.6 mm (28 mesh) (0.024 in.)
for acid leaching, less than 0.7 mm (200 mesh) for alkaline
leaching in rod or ball mills using water (or preferably, leach)
to obtain a pulp density of about two-thirds solids. Screw
classifiers, thickeners, or cyclones are sometimes used to
control size or pulp density.
Acid Leach. Ores with a calcium carbonate (CaC03_) content of
less than 12 percent are preferentially leached in sulfuric acid,
which extracts values quickly (in four hours to a day), and at a
lower capital and energy cost than an alkaline leach. Any
tetravalent uranium must be oxidized to the uranyl form by adding
an oxidizing agent (typically, sodium chlorate or manganese
dioxide), which is believed to facilitate the oxidation of U(4)
to U(6) in conjunction with the reduction of Fe (3) to Fe (2) at
a redox (reduction/oxidation) potential of about minus 450 mV.
Free-acid concentration is held to between 1 and 100 grams per
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liter. The larger concentrations are suitable when vanadium is
to be extracted. The reactions taking place in acid oxidation
and leaching are:
2U02 + 02 2U03_
2U03_ + 2H2SO£ + 5H20 2(U02S04)
. 7H20
Uranyl sulfate (U02S04_) forms a complex, hydrouranyl trisulfuric
acid (H4U02_(S04)3_ in the leach, and the anions of this acid are
extracted for value.
Alkaline Leach. A solution of sodium carbonate (40 to 50 g per
liter) in an oxidizing environment selectively leaches uranium
and vanadium values from their ores. The values may be
precipitated directly from the leach by raising the pH and adding
sodium hydroxide. The supernatant can be recycled after its
exposure to carbon dioxide. A controlled amount of sodium
bicarbonate (10 to 20 g per liter) is added to the leach to lower
pH which prevents spontaneous precipitation.
This leaching process is slower than acid leaching since other
ore components are not attached and these ore components tend to
shield the uranium values. Therefore, alkaline leach is used at
elevated temperatures of 80 to 100 degrees Celsius (176 to 212
degrees Fahrenheit) and is subjected to the hydrostatic pressure
at the bottom of a 15 to 20 m (49.2 to 65.6 ft) tall tank which
contains a central airlift for agitation (Figure A-18). In some
mills, the leach tanks are pressurized with oxygen to increase
the rate of reaction which normally takes one to three days. The
alkaline leach process is characterized by the following
reactions:
2U02 + 02 2U03 (oxidation)
3Na2_(C03_) + U03_ + H2_0 2NaOH +
Na4(U02_) (CO3_)3_ (leaching)
2NaOH + C02. Na2C03_ + H2_0
(recarbonization)
2Na4(U02) (C03_)3_ + 6NaOH
Na2U207_ + 6Na2_C03_ + 3H2_0 (precipitation)
Alkaline leaching can be applied to a greater variety of ores
than is currently being done; however, this process, because of
its slowness, apparently involves greater capital expenditures
per unit production. In addition, the purification of yellow-
cake, generated in a loop using sodium as the alkali element,
consumes an increment of chemicals that tend to appear in stored
or discharged wastewater. Purification to remove sodium ion is
necessary both to meet the specifications of American uranium
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processors and for the preparation of natural uranium dioxide
fuel. The latter process will be used to illustrate the problem
caused by excess sodium. Sodium diuranate may be considered as a
mixture of sodium and uranyl oxides—i.e., Na2JJ20^7 = Na20 + 2U03_.
The process of generating UO2^ fuel pellets from a yellowcake feed
involves reduction by gaseous ammonia at a temperature of a few
hundred degrees C. At this temperature, ammonia thermally
decomposes into hydrogen, which reduces the U03_ component to U02^
and nitrogen (which acts as an inert gas and reduces the risk of
explosion in and around the reducing furnace). With sodium
diuranate as a feed, 'the process results in a mix of U02^ and Na2_0
that is difficult to purify (by water leaching of NaOH) without
impairing the ceramic qualities of uranium dioxide. When, in
contrast, ammonium diuranate is used as the feed, all byproducts
are gaseous, and pure U02^ remains. The structural integrity of
this ceramic is immediately adequate for extended use in the
popular CANDU (Canadium deuterium-uranium) reactors. Sodium ion,
as well as vanadium values, can be removed from raw yellowcake
(sodium diuranate) produced by alkaline leaching. First, the
yellowcake is roasted, and some of the sodium ion forms water-
soluble sodium vanadate, while organics are carbonized and burned
off. The roasted product is water leached, yielding a V205
concentrate as described below. The remaining sodium diuranate
is redissolved in sulfuric acid,
Na2_U201 + 3H2S04 Na2S04_ +
3H20 + 2(U02_)SOi
and the uranium values are precipitated with ammonia and filtered
to yield a yellowcake (ammonium diuranate or UCB) that is low
in sodium.
U02S04 + H2O + 2NH3.
(NH4)2S04 + U03
The byproduct that is formed, sodium sulfate, being classed
approximately in the same pollutant category as sodium chloride,
requires expensive treatment for its removal. Ammonium-ion dis-
charges, which might result from an ammonium carbonate leaching
circuit, are viewed with more concern, even though there is a
demand for ammonium sulfate for fertilization of alkaline south-
western soils. Ammonium sulfate could be generated by neutraliz-
ing the wastes of the ammonium loop with sulfuric acid wastes
from acid leaching wastes. Opponents of a tested ammonium pro-
cess argue that nitrites, an intermediate oxidation product of
accidentally discharged ammonium ion, present a present health
hazard more severe than from sulfate ion.
Vanadium Recovery. Vanadium, found in carnotite (K 2_ (UO 2_) 2 (V04_) 2_
3H20) as well as in heavy metal vanadates—e.g., vanadinite
(9PbO . 3V2_05_ . PbCl) — is converted to sodium orthovanadate
(Na3V04_), which is water-soluble, by roasting with sodium
575
-------
chloride or soda ash (Na2_03_). After water leaching, ammonium
chloride is added, and poorly soluble ammonium vanadates are
precipitated:
Na3V04 + 3NH4C1 + H20 SNaOH +
NH4V03_ + 2NH40H
(ammonium metavanadate)
Na3_V04 + 3NH4C1 3NaCl +
(NH4_)3_V04
(ammonium orthovanadate)
The ammonium vanadates are thermally decomposed to yield vanadium
pentoxide:
3(NH4)3_V04 — 6NH3 + 3H20 +
V205
A significant fraction (86 to 87 percent) of V2_05^ is used in the
ferroalloys industry. There, ferrovanadium has been produced in
electric furnaces (the following reaction applies):
V5. + Fe203_ + 8C SCO + 2FeV
or by aluminothermic reduction (See Glossary) in the presence of
scrap iron.
Air pollution problems associated with the salt roasting process
have led many operators to utilize a hydrometallurgical process
for vanadium recovery which is quite similar to uranium recovery
by acid leaching and solvent exchange. The remainder of V2O5^
production is used in the inorganic chemical industry.
Concentration and Precipitation. Approximately one metric ton of
ore with a grade of about 0.2 percent is treated with one metric
ton (or cubic meter) of leach, and the concentration(s) of
uranium and/or vanadium in the pregnant solution are also about
0.2 percent. If values were directly precipitated from the
solution, a significant fraction of the values would remain in
solution. Therefore yellowcake is recycled and dissolved in a
pregnant solution to increase precipitation yield. Direct
precipitation by raising the pH is effective only for an alkaline
leach, because it is more selective for uranium and vanadium. If
this technique were applied to the acid leach process, most heavy
metals—particularly, iron—would be precipitated, thus severely
contaminating the product.
Uranium (or vanadium and molybdenum) in the pregnant leach liquor
can be concentrated through ion exchange or solvent extraction.
Typical concentrations in the eluate of some of the processes are
shown in Table A-3.
576
-------
Precipitation of uranium from the eluates is achievable without
recycling yellowcake, and the selectivity of these processes
under regulated conditions (particularly, pH), improves the
purity of the product.
All concentration processes operate best in the absence of
suspended solids, and considerable effort is made to reduce the
solids content of pregnant leach liquors (Figure A-19-a). A dis-
tinction is made between quickly settling sands that are not
tolerated in any concentration process and slimes that can be
accomodated to some extent in the resin-in-pulp process (Figure
A-19-b-c). Sands are often repulped, by the addition of some
wastewater stream, to facilitate flow to the tailing pond.
Consequently, there is some latitude for the selection of the
wastewater sent to the tailing pond, and mill operators can take
advantage of this fact in selecting environmentally sound waste
disposal procedures.
Ion exchange and solvent extraction (Figure A-19-b-e) are based
on the same principle: Polar organic molecules tend to exchange
a mobile ion in their structure—typically, C1-, N03/-, HS04_-,C03_
(anions), or H+ or Na+ (cations)—for an ion with a greater
charge or a smaller ionic radius. For example, let R be the
remainder of the polar molecule (in the case of a solvent) or
polymer (for a resin), and let X be the mobile ion. Then, the
exchange reaction for the uranyltrisulfate complex is:
4RX + (U02(S04)3.)
R4U02(S04)3. + 4X
This reaction proceeds from left to right in the loading process.
Typical resins adsorb about ten percent of their mass in uranium
and increase by about ten percent in density. In a concentrated
solution of the mobile ion—for example, in N-hydrochloric acid—
the reaction can be reversed and the uranium values are eluted —
in this example, as hydrouranyl trisulfuric acid. In general,
the affinity of cation exchange resins for a metallic cation
increases with increasing valence (Cr+++, Mg-n-, Na+), and because
of decreasing ionic radius, with increasing atomic number (92U,
42Mo, 23V). The separation of hexavalent 92U cations by IX or SX
should prove to be easier than that of any other naturally
occurring element.
Uranium, vanadium, and molybdenum—the latter being a common ore
consitutent—almost always appear in aqueous solutions as
oxidized ions (uranyl, vanadyl, or molybdate radicals). Uranium
and vanadium also combine with anionic radicals to form trisul-
fates or tricarbonates in the leach. The complexes react anion-
ically, and the affinity of exchange resins and solvents is not
simply related to fundamental properties of the heavy metal
(uranium, vanadium, or molybdenum), as is the case in cationic
exchange reactions. Secondary properties, including pH and redox
potential, of the pregnant solutions influence the adsorption of
577
-------
heavy metals. For example, seven times more vanadium than
uranium is adsorbed on one resin at pH 9 whereas at pH 11, the
ratio is reversed, with 33 times as much uranium as vanadium
being captured. These variations in affinity, multiple columns,
and control of variations in affinity, multiple columns, and con-
trol of leaching time with respect to breakthrough (the time when
the interface between loaded and regenerated resin, e.g., Figure
A-19-d, arrives at the end of the column) are used to make an IX
process specific for the desired product.
In the case of solvent exchange, the type of polar solvent and
its concentration in a typically nonpolar diluent (e.g., kero-
sene) effect separation of the desired product. The ease with
which the solvent is handled (Figure A-19-e) permits the con-
struction of multistage co-current and countercurrent SX concen-
trators that are useful even when each stage effects only partial
separation of a value from an interferent. Unfortunately, the
solvents are easily polluted by slimes, and complete liquid/solid
separation is necessary. IX and SX circuits can be combined to
take advantage of both the slime resistance of resin-in-pulp ion
exchange and the separatory efficiency of solvent exchange (Eluex
process-Figure A-19-f). The uranium values are precipitated with
a base or a combination of base and hydrogen peroxide. Ammonia
is preferred by a plurality of mills because it results in a
superior product, as mentioned in the discussion of alkaline
leaching. Sodium hydroxide, magnesium hydroxide, or partial
neutralization with calcium hydroxide followed by magnesium
hydroxide precipitation, are also used. The product is rinsed
with water that is recycled into the process to preserve values,
then filtered, dried and packed into 200-liter (55-gallon) drums.
The strength of these drums limits their capacity to 450 kg (1000
pounds) of yellowcake which occupies 28 percent of the drum
volume.
Figure A-19-g illustrates the Split Elution Concentration
process. Figure A-20 illustrates a "Generalized Flow Diagram for
Production of Uranium, Vanadium and Radium."
Antimony Ore Mil 1 ing Processes
Antimony is recovered from antimony ore and as a byproduct from
silver and lead concentrates.
Only a small percentage of antimony (13 percent in 1972) is
recovered from ore being mined primarily for its antimony
content. Nearly all of this production can be attributed to a
single operation which is using a froth flotation process to
concentrate stibnite (Sb2S3.) (Figure A-21).
The bulk of domestic production of antimony is recovered as a
byproduct of silver mining operations in the Coeur d'Alene dis-
trict of Idaho. Antimony is present in the silver-containing
mineral tetrahedrite and is recovered from tetrahedrite concen-
578
-------
trates in an electrolytic antimony extraction plant owned and
operated by one of the silver mining companies in the Coeur
d'Alene district. Mills are usually penalized for the antimony
content in their concentrates. Therefore, the removal of anti-
mony from the tetrahedrite concentrates not only increases their
value, but the antimony itself then becomes a marketableiitem.
Antimony is also contained in lead concentrates and is ultimately.
recovered as a byproduct at lead smelters—usually as antimonial
lead. This source of antimony represents about 30 to 50 percent
of domestic production in recent years.
Titanium Ore Milling Processes
The method of mining and beneficiating titanium minerals depends
upon whether the ore is contained ina sand or rock deposit. Sand
deposits occurring in Florida, Georgia, and New Jerrsey contain 1
to 5 percent Ti02% and are mined with floating suction or bucket-
line dredges handling up to 1,088 metric tons (1,200 short tons)
of material per hour. The sand is treated by wet graiity methods
using spirals, cones, sluices, or jigs to produce a bulk, mixed,
heavy-mineral concentra As many as five individual marketable
minerals are then separated from the bulk concentrate by a
cbination of dry separation techniques using magnetic and
electrostatic (high-tension) separators, sometimes in conjunction
with dry and wet gravity concentrating equipment.
High-tension (HT) electrostatic separators are employed to
separate the titanium minerals from the silicate minerals-. The
minerals are fed onto a high-speed spinning rotor, and a heavy
corona (glow given off by a high voltage charge) discharge is
aimed toward the minerals at the point where they would normally
leave the rotor. The minerals of relatively poor electrical con-
ductance are pinned to the rotor by the high surface charge they
recieve on passing through the high voltage corona. The minerals
of relatively high conductivity do not readily hold this surface
charge and so leave the rotor in their normal trajectory. Titan-
ium minerals are the only ones present of relatively high elec-
trical conductivity and are, therefore, thrown off the rotor.
The silicates are pinned to the rotor and are removed by a fixed
brush.
Titanium minerals undergo final separation in induced-roll
magnetic separators to produce three products: ilmenite,
leucoxine, and rutile. The separation of these minerals is based
on their relative magnetic properties which, in turn, are based
on their relative iron content: ilmenite has 37 to 65 percent
iron, leucoxine has 30 to 40 percent iron, and rutile has 4 to 10
percent iron.
Tailings from the HT separators (nonconductors) may contain
zircon and monazite (a rare-earth mineral). These heavy minerals
are separated from the other nonconductors (silicates) by various
579
-------
wet gravity methods (i.e., spirals or tables). The zircon (non-
magnetic) and monazite (slightly magnetic) are separated from one
another in induced-roll magnetic separators.
Beneficiation of titanium minerals from beach-sand deposits is
illustrated in Figure A-22.
Ilmenite is also currently mined from a rock deposit in New York
by conventional open-pit methods. This ilmenite/magnetite ore,
averaging 18 percent Ti02^ is crushed and ground to a small
particle size. The ilmenite and magnetite fractions are
separated in a magnetic separator, the magnetite being more
magnetic due to its greater iron content. The ilmenite sands are
further upgraded in a flotation circuit. Beneficiation of
titanium from a rock deposit is illustrated in Figure A-23.
580
-------
Table A-1
REAGENTS USED FOR FLOTATION OF IRON ORES
{flngmt quantities represent approximate maximum usage*. Exact chemical composition of reagent
may be unknown.)
1. Anionic Flotation of Iron Oxides (from crude ore)
Petroleum sulfonata: 0.5 kg/metric ton (1 Ib/short ton)
Low-rosin, tall oil fatty acid: 0.25 kg/matric ton (0.5 Ib/short ton)
Suifuric acid: 1.25 kg/metric ton (2.5 Ib/short ton) to pH3
No. 2 fuel oil: 0.15 kg/metric ton (0.3 Ib/short ton)
Sodium silicate: 0.5 kg/metric ton (1 Ib/short ton)
2. Anionic Flotation of Iron Oxides (from crude ore)
Low-rosin tall oil fatty acid: 0.5 kg/metric ton (1 Ib/short ton)
3. Cationic Rotation of Hematite (from crude ore)
Rosin amine acetate: 0.2 kg/metric ton (0.4 Ib/short ton)
Suifuric acid: 0.15 kg/metric ton (0.3 Ib/short ton)
Sodium fluoride: 0.15 kg/metric ton (0.3 Ib/short ton)
(Plant also includes phosphate flotation and pyrite flotation steps. Phosphate flotation employs
sodium hydroxide, tall oil fatty acid, fuel oil, and sodium silicate. Pyrita flotation employs
xanthata collector.)
4. Cationic Flotation of Silica (from crude ore)
Amine: 0.15 kg/metric ton (0.3 Ib/short ton)
Gum or starch (tapioca fluor): 0.5 kg/metric ton (1 Ib/short ton)
Methyfisooutyl caroinol: as required
5. Cationic Flotation of Silica (from magnetite concentrate)
Amine: 5 g/metric ton (0.01 Ib/short ton)
Methylisobutyl carfainol: as required
581
-------
Table A-2
VARIOUS FLOTATION METHODS AVAILABLE FOR PRODUCTION
OF HIGH-GRADE IRON-ORE CONCENTRATE
1. Anionic flotation of specular hematite
2. Upgrading of natural magnetite concentrate by cationic flotation
3. Upgrading of artificial magnetite concentrate by cationic flotation
4. Cationic flotation of crude magnetite
5. Anionic flotation of silica from natural hematite
6. Cationic flotation of silica from non-magnetic iron formation
582
-------
Table A-3
URANIUM CONCENTRATION IN IX/SX ELUATES
PROCESS
U3Og CONCENTRATION (%)
Ion exchange
Resin-in-pulp
Fixed-bed IX:
Chloride elution
Nitrate elution
Moving-bed IX:
Nitrate elution
0.5 to 1.0
1.0 to 2.0
1.9
Solvent extraction
Alky) phosphates, HCI eluent
Amex process
Dapex process
Split eiution minewater treatment
20.0 to 60.0
3 to 4
5.0 to 6.5
1.2to 1.6
IX/SX combination
Eluex process
3.0 to 7.5
583
-------
ORE
CRUSHING AND
SCREENING
f
BLENDING
1
T
CONCENTRATING PROCESSES:
PHYSICAL
1 1
T T
WASHING JIGG
SINTERING
1 i
|W_ MAGNETIC ^£1
ING SEPARATION MED
itfAKAriON SEPARA
1
AGGLOMERATION
PROCESSES
PELLETIZING NODULIZINC
CHEMICAL
|
/Y-
IA FLOTATION
TION
j
BRIQUETTING
T
t
TO STOCK PILE AND/OR SHIPPING
Figure A-1
BENEFICIATION OF IRON ORES
584
-------
DENSIFYING THICKENER
UNDERFLOW
CONDITIONERS
ROUGHER FLOTATION
ROUGHER
TAIL
TO
TAILING
BASIN
ROUGHER
CONCENTRATE
(10 CELLS)
FROTH OF
FIRST 2 CELLS
I
CLEANER FLOTATION
CLEANER
TAIL
CLEANER
CONCENTRATE
(8 CELLS)
FROTH OF
FIRST 2 CELLS
I
I
RECLEANER FLOTATION
RECLEANER
TAIL
I
RECLEANER
CONCENTRATE
(7 CELLS)
I
TOTAL
FLOTATION
CONCENTRATE
j
TO AGGLOMERATION
(FIGURE 111-4)
Figure A-2
IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
585
-------
CRUSHED CRUDE ORE
i
)M
I
COBBER MAGNETIC SEPARATION
CONCENTRATE
J
BALL MILL
CLEANER MAGNETIC SEPARATION
CONCENTRATE
JHYDROCYCLONE
OVERSIZE UNDERS1ZE
i
i
FINISHER MAGNETIC SEPARATION
CONCENTRATE
THICKENING
I
LTI
T
HYDROSEPARATOR
i
JCENTRATE
TAILING
TO TAILING BASIN
TO PELLETIZ1NG
(FIGURE MM)
Figure A-3
MAGNETIC TACONITE BENEFICIATION FLOWSHEET
586
-------
o
(— I
Ul
x
-------
CONCENTRATE FILTER CAKE
BALLING DRUM
I
SCREEN
UNDERSIZE OVERSIZE
I
AGGLOMERATION FURNACE
PELLETS
t
EXHAUST GASES
t
TO STOCK PILE TO ATMOSPHERE
AND/OR SHIPPING
Figure A-5
AGGLOMERATION FLOWSHEET
588
-------
ORE K01XC«I
OREI014)4XCul -ORE ^ c?ne—CE1A«C
ORE WASTE DUMPS
WASTE 1 '"'MARY - SCREENIIgG
OOMP «*' "*,SIJU 1 CRUSHER ~ SCREENING
(ACIOI I*C.O> IAC.DI M,,fn I
. , ^ r , _ -_,. OXIOf S'HF'Of SECONDARY
ORE CRUSHER
ACID ACID ACID I
SOLUTION SOL N SOL N J
ACIO ACID ACIO
RECYCLED RECYCLED RECYCLED SCREENIW3
1
PRECp''!T*T'ON HEAP _ TERT.ARY
" |*^ CRUSHER 'MIXED'
,; " ACIO ' i ' iULti°
SOLUTION 1 __.« . I. . ,,.. -,
AC1UHH-YCUU . 1— | CYCLONES |
' SPONfiE IRON f
CEMENT , COPPER AND BALL ^. T... ..,„
COPPER ., , "A"" IRON MILL •»-|w*r*>1
1 , i 1 IRE Ait
TflSMEIT(R TAILING — ,...r* ....... _ FLOTATION •* '
| POND THICKCNCII.. -* ^^^ ^
\ \ \
HfHNERy RECYCLED WATER CONCENTHATE
L ' J '* L
REflNESY
TO SMELTER
CEMENT
COPPER
t
PRECIflTATIO*
PLANT
OXIDE/
E ORE
t
i
WASH
WATER
. VAT LEACH
NTS! IACIOI
.
TAILS
BARREN
P
SI
i r
ELECTRO
WINNING
FACILITY
AND
OW
i ^uiFinc
[ HITCH*") 1
T TO DUMP
BYPRODUCT COI'I'tHISI
IWOLYBDENUM CUNCENTKATE
T to iMELTEH
1U SMUIIH 1
I *
CATHODE
COPPER
1
TO MA
IOR REF
REGNANT
DLUTION
BXf T
INCHYI
HE FIWEMY
MAHKET
Figure A-6
GENERAL FLOW DIAGRAM DEPICTING METHODS FOR
TYPICAL RECOVERY OF COPPER FROM ORE
589
-------
0)
n
3
ofl
Q H
O M
590
-------
ORE MINING DRAINAGE
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS •
FUELS
LUBRICANTS
TO POND
• AND/OH
MILL
WATER FROM MINE,
RECYCLE OR OTHER
REAGENTS
CONCENTRATE
FINAL LEAD
CONCENTRATE
THICKENING
AND FILTRATION
USUALLY
RECYCLED
TO PROCESS
WATER SYSTEM
ZINC ROUGHER
CONCENTRATE
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS
EXCESS REAGENTS
'"I
TO SUBSURFACE
DRAINAGE
CONCENTRATE
TO ZINC
SMELTER
USUALLY RECYCLED
TO PROCESS
WATER SYSTEM
Figure A-8
TYPICAL LEAD-ZINC MINING AND PROCESSING OPERATION
591
-------
out
TO MIlTIft
Figure A-9
CYANIDATION OF GOLD ORE: VAT LEACHING OF SANDS
AND 'CARBON-IN-PULP' PROCESSING OF SLIMES
592
-------
ORE
i
CRUSHING
GRINDING
CONDITIONING
COUNTERCURRENT
LEACHING IN
THICKENERS
I
PRECIPITATION
OF GOLD FROM
LEACHATE BY
ADDITION OF
ZINC DUST
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
PRECIPITATE
FILTERED AND
THICKENED
TO SMELTER
REAGENTS (CN)
BARREN
PULP
TAILJNG-POND
DECANT
RECYCLED
BARREN SOLUTION
RECYCLED
Figure A-10
CYANIDATTON OF GOLD ORE: AGITATION/LEACH PROCESS
593
-------
ORE
CRUSHING
GRINDING
CONDITIONING
1
SELECTIVE i
FROTH
FLOTATION '>
CONCENTRATE
FILTERED
AND THICKENED
TO SMELTER
FLOTATION CIRCUIT
TAILINGS
REAGENTS (CN)
LEACHING IN
THICKNERS
I
.BARREN.
PULP
TO TAILING
'POND
PRECIPITATION OF GOLD
FROM LEACHATE BY
ADDITION OF ZINC DUST
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
PRECIPITATE FILTERED
AND THICKENED
TO SMELTER
Figure A-11
FLOTATION OF GOLD-CONTAINING MINERALS WITH RECOVERY OF
RESIDUAL GOLD VALUES BY CYANIDATION
594
-------
ORE
AGENTS [• *--<—
_ 1 T
NO. 1
FLOTATION CIRCUIT
NO. 2
FLOTATION CIRCUIT
RETREATMENT
CIRCUIT
PYRITE
CONCENTRATE
NO. 3
FLOTATION CIRCUIT
FINAL Ag
CONCENTRATE'
FINAL
TAILINGS
•CONTAINS
25.7 TO 44.6 KILOGRAMS PER
METRIC TON
(750-1300 OUNCES PER SHORT TON):
25 TO 32% COPPER
0 TO 18% ANTIMONY
FINAL PYRITE.
CONCENTRATET
CONTAINS 3.43 KILOGRAMS PER
METRIC TON (100 TROY OUNCES
PER SHORT TON)
Figure A-12
RECOVERY OF SILVER ORE BY FROTH FLOTATION
595
-------
MINING 1
ORE
t
CRUSMINGS.
WEIGHING. AND
SCREENING
[
r^
BALL
MILLS
1 '
CYCLONES
l
UNDERFLOW
OVERFLOW
TAILS
ROUGHER FLOAT
- CONCENTRATE--
— MIDDLINGS —
MIDDLINGS
__i—
SCAVENGER
FLOAT
14 STAGES WITH
REGRIND AND
INTERNAL RECYCLE!
T ~~
CONCENTRATE -
MIDDLINGS •
CLEANER
FLOAT
16 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
TAILS
CONCENTRATE
-UNDERFLOW —
-UNDERFLOW
OVERFlOW
-(
CYCLONES
>
UNDERFLOW-
-. OVERFLOW
-THICKENER RECLAIM —
J WATER
[DRYER J
MOLYBDENUM
PRODUCT
TO TAILING
POND
Figure A-1 3
SIMPLIFIED MOLYBDENUM MILL FLOWSHEET
SHOWING RECOVERY OF MOLYBDENITE ONLY
596
-------
CONCENTRATE
LIGHT TO TAILS
LIGHT TO TAILS
MONAZITE
CONCENTRATE —
TO TAILS
CRUSHING
(3 STAGES)
1
28% + 3 MESH
*
GRINDING
BALL MILLS
1
36% + 100 MESH
>|
t
FLOTATION
t
FLOTATION
96% OF MILL FEED
t
GRAVITY
HUMPHREY'S SPIRALS
1
PYRITE
FLOTATION
1
TAILS
t
TABLES
~T
MONAZSTE j
FLOTATION ,
— =f=^
MAGNETIC 1
SEPARATION |
I
NONMAGNETIC
TIN CONCENTRATE , >
CONCENTRATE
CLEANER
FLOTATION
{4 STAGES)
\ '
DRYING
I
MOLYBDENUM
CONCENTRATE
(93% + MoS2)
-TAILINGS
MAGNETIC TUNGSTEN
CONCENTRATE
Figure A-14
SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM
SHOWING RECOVERY OF MOLYBDENITE AND BYPRODUCTS
597
-------
ORE
SULFIDE
FLOTATION
CYCLONE
25%.
SLIMES
75% SANDS
i
GRAVITY
TABLES
TAILINGS
OVERFLOW
THICKENER
SCHEELITE
FLOTATION
HCI LEACH
(15 TO 20% OF
FRACTION)
TUNGSTEN
CONCENTRATE
Figure A-1 5
SIMPLIFIED FLOW DIAGRAM FOR A SMALL
TUNGSTEN CONCENTRATOR
598
-------
6-10%
NaCL
H2S04
TERTIARY
AMINES
1.5 - 2.0% V2O5
t
GRINDING
PELLETIZING
ROASTING
850°C (1562°F)
NaVO,
LEACHING AND
ACIDIFICATION
pH 2.5 - 3.5
(Na6V10°28
SODIUM DECAVANADATE)
SOLVENT EXTRACTION
NH4OH
PRECIPITATION
1
NH4V03
PRECIPITATE
CALCINING
T
V205 PRODUCT
Figure A-16
ARKANSAS VANADIUM PROCESS FLOWSHEET
599
-------
ORE
CLAY HOPPER
AUTOGENOUS
MILL
CLASSIFIER
I
ROUGHER
FLOTATION CELLS
CLEANER
FLOTATION CELLS
I
CLEANER
FLOTATION CELLS
i
THICKENER
I
FILTER
I
CONCENTRATE
PRODUCT
OVERFLOW
WATER
TAILINGS
Figure A-17
FLOW DIAGRAM FOR BENEFICIATION OF
MERCURY ORE BY FLOTATION
600
-------
VENT
..:v-
.•:f*."*~".-;-S):--f.
W$-'W£.':
?:°:^-3?i
:**•>.•••. :v..o"
••'v.1; .?•.<* : ;
•. S-« .—.•,>•-*••
"»i:.> °^ '-
< •• • .• .- • Q •
v^y^J?
y'ix'&i-
&
,^>''. :/0 • o '«•
^•5v';v-:-"-
«-f.--.r;>..^:;. -
.^* ••••>'••- - •
•.••rt?^r?";-*'
^%4^.--
-^f^^>
x\
, LEACH
AIRLIFT
AIR COMPRESSOR
Figure A-18
PACHUCA TANK FOR ALKALINE LEACHING
601
-------
FROM
LEACH
PREGNANT
LEACH LIQUOR
SLIMES
SLIMY PULP TO
RESIN-IN-PULPIX
CLEAR LEACH LIQUOR
TO COLUMN IX OR SX
a) LIQUID/SOLID SEPARATION
SAND
TAILINGS
SLIMY,
PREGNANT—?
PULP
RESIN IN OSCILLATING BASKET
b) RESIN-IN-PULPPROCSSS: LOADING
BARREN
•;> PULP
TO TAILINGS
BARREN
ELUANT
PREGNANT
ELUATS TO
PRECIPITATION
c) RESIN-IN-PULP PROCESS: ELUT1NG
Figure A-19
CONCENTRATION PROCESSES AND TERMINOLOGY
602
-------
BARREN ELUANT
ELUTED (OR
REGENERATED)
RESIN
LOADED
RESIN
PREGNANT ELUATE
TO PRECIPITATION
d) FIXED-BED COLUMN ION EXCHANGE/ELUT1ON
PREGNANT
LEACH
LIQUOR
o o*
LOADING
LEACH
SOLVENT
LOADED
ORGANIC
iui
BARREN STRIPPED
A ELUANT SOLVENT
BARREN
LJQUOR
T
PHASE
SEPARATION STRIPPING
e> SOLVENT EXTRACTION
RECYCLE
ELUANT
SOLVENT
IX
ti-UAN f
~\ s~ ^
1 1
/ \
IX
PREGNANT 1
ELUATE II1
PHASE
SEPARATION
PREGNANT
ELUATE
PARTIALLY STORAGE)
STRIPPED \ , LOADED
RESIN X _ ' RESIN
9) SPLIT ELUT1ON
PRECIPITATION
f) ELUEX PROCESS
Figure A-19
CONCENTRATION PROCESSES AND TERMINOLOGY (Continued)
603
-------
MINING
I
ORE TREATMENT
LEACHING
'-1
LIQUID/SOLID
SEPARATION
I"
I
ION EXCHANGE
SOLVENT EXTRACTION
PATH I
I
PATHIH
PATH It
PRECIPITATION
TO
STOCKPILE
URANIUM
CONCENTRATE
I
VANADIUM
BYPRODUCT
RECOVERY
TO
STOCKPIL
Figure A-20
GENERALIZED FLOW DIAGRAM FOR PRODUCTION OR URANIUM
VANADIUM, AND RADIUM
604
-------
MINING
ORE
i
CRUSHING
r
GRINDING
i
ROUGHER
FLOTATION
I
FROTH
-TAILS
I
CLASSIFICATION
•TAILS-
SCAVENGER
FLOTATION
CLEANER
FLOTATION
I
FROTH
I
TO
WASTE
FROTH
FILTER
FILTRATE
I
THICKENER
I
WASTE
FINAL
CONCENTRATE
T
TO SHIPPING
Figure A-21
BENEFICIATION OF ANTIMONY SULFIDE ORE BY FLOTATION
605
-------
I WET MILL
| OR1 MILL
Figure A-2 2
BENEFICIATION OF HEAVY MINERAL BEACH SANDS
606
-------
MINING
I
ORE
CRUSHING
GRINDING
I
CLASSIFICATION
MAGNETIC
SEPARATION
MAGNETICS-
I
NONMAGNETICS
MAGNETITE
)
r
ILMENITE
AND GANGUE
DEWATERER
i
FLOTATION
CIRCUIT
T
i
THICKENER
TO
WASTE
1
FILTER
I
DRIER
CONCENTRATE
J
TO SHIPPING
Figure A-23
BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
607
-------
APPENDIX B
PRELIMINARY INTERIM PROCEDURE
FOR
FIBROUS ASBESTOS
608
-------
JAN
DRAFT
INTERIM METHOD
FOR ASBESTOS
IN WATER
by
Charles H. Anderson and J. MacArthur Long
Revised December 29, 1973
Analytical Chemistry 3ranch
o. Environmental Protection Agency
Environmental Research Laboratory
College Station Road
Athens, Georgia J0605
609
-------
Preface to Revised EPA Interim Method for
Determining Asbestos in water
In July 1976 the Preliminary Interim Method for Determining
Asbestos in Water was issued by the Athens Environmental
Research Laboratory. That method was perceived as representing
the current state-of-the-art in asbestos analytical method-
ology. The objective of writing the method was to present a
procedure analytical laboratories could follow that would result
in a better agreement of analytical results. In the past two
years, a significant amount of additional experimental work has
generated data that provide the basis for a more definitive
method than was possible previously.
This revised Interim Method reflects the improvements that have
been made in asbestos analytical methodology since the initial
procedure was drafted. The general approach to the analytical
determination, however, remains the same as previously out-
lined. That is, asbestos fibers are separated from water by
filtration on a sub-micron pore size membrane filter. The
asbestos fibers are then counted, after dissolving the filter
material, by direct observation in a transmission electron
microscope.
The major change in the initial procedure is the elimination of
the condensation washer as a means of sample preparation.
Intra- and inter-laboratory precision data for the method are
presented. Also, a suggested statistical evaluation of grid
fiber counts is included.
DRAFT
610
-------
ASBESTOS
(Interim Method)
(Transmission Electron Microscopy Method)
1. Scope and Application
1.1 This method is applicable to drinking water and
water supplies.
1.2 The method determines the number of asbestos fibers/
liter, their size (length and width), the size
distribution, and total mass. The method distin-
guishes chrysotile from amphibole asbestos. The
detection limits are variable and depend upon the
amount of total extraneous particulate matter in the
sample as well as the contamination level in the
laboratory environment. Under favorable circum-
stances 0.1 MFL (million fibers per liter) can be
detected. The detection limit for total mass of
asbestos fibers is also ^variable and depends upon
the fiber size and size distribution in addition to
the factors affecting the total fiber count. The
detection limit under favorable conditions is in the
order of 0.1 ng/1.
1.3 The method is not intended to furnish a complete
characterization of all the fibers in water.
1.4 it is beyond the scope of this method to furnish
detailed instruction in electron microscopy,
electron diffraction or crystallography. It is
assumed that those using this method will be suffi-
ciently knowledgeable in these fields to understand
the methodology involved.
2. Summary of Method
2.1 A variable, known volume of water sample is filtered
through a membrane filter of sufficiently small pore
size to trap asbestos fibers. A small portion of
the filter with deposited fibers is placed on an
electron microscope grid and the filter material
removed by gentle solution in organic solvent. The
material remaining on the electron microscope grid
is examined in a transmission microscope at high
magnification. The asbestos fibers are identified
by their morphology and electron diffraction pattern
and their length and width are measured. The total
area examined in the electron microscope is deter-
mined and the number of asbestos fibers in this area
-------
is counted. The concentration in MFL (millions of
fibers/liter) is calculated from the number of
fibers counted, the amount of water filtered, and
the ratio of the total filtered area/sampled filter
area. The mass/liter is calculated from the assumed
density and the volume of the fibers.
3. Definitions
Asbestos - A generic term applied to a variety of
commercially useful fibrous silicate minerals of the
serpentine or amphibole mineral groups.
Fiber - Any particulate that has parallel sides and a
length/width ratio greater than or equal to 3:1.
Aspect Ratio - The ratio of length to width.
Chrysotile - A nearly pure hydrated magnesium silicate, the
fibrous form of the mineral serpentine, possessing a
unique layered structure in which the layers are
wrapped in a helical cylindrical manner about the
fiber axis.
Amphibole Asbestos - A double chain fibrous silicate
mineral consisting of 814011 units, laterally
linked by various cations such as aluminum, calcium,
iron, magnesium, and sodium. The members of the
amphibole asbestos consist of the following:
crocidolite, cummingtonite-gruenerite, and the
fibrous forms of tremolite, actinolite and antho-
phyllite. These minerals consist of or contain
fibers formed through natural growth processes.
Mineral fragments that: conform to the definition of
a fiber and that are formed through a crushing and
milling process are analytically indistinguishable
from the naturally formed fibers by this method.
Detection Limit - The calculated concentration in MFL,
equivalant to one fiber above the background or
blank count. (Section 8.6).
Statistically Significant - Any concentration based upon a
total fiber count of five or more in 20 grid squares.
4. Sample Handling and Preservation
4.A Sampling
It is beyond the scope of this procedure to furnish
detailed instructions for field sampling; the general
principles ot sampling waters are applicable. There are
some considerations that apply to asbestos fibers, a
special type ot particulate matter. These fibers are
612
-------
small, and in water range in length from . 1 urn to 20 ^m or
more. Because of the range of size there may be a verti-
cal distribution of particle sizes. This distribution
will vary with depth depending upon the vertical distri-
bution of temperature as well as the local meteorological
conditions. Sampling should take place according to the
objective of the analysis. If a representative sample of
a water supply is required a carefully designated set of
samples should be taken representing the vertical as well
as the horizontal distribution and these samples compos-
ited for analysis.
4.1 Containment Vessel
The sampling container shall be a clean conventional
polyethylene, screw-capped bottle capable of holding
at least one liter. The bottle should be rinsed at
least two times with the water that is being sampled
prior to sampling.
NOTE: Glass vessels are not suitable as sampling
containers.
4.2 Quantity of Sample
A minimum of approximately one liter of water is
required and the sampling container should not be
filled. It is desirable to obtain two samples from
one location.
4.3 Sample Preservation
No preservacives should bemadded during sampling and
the addition of acids should be particularly
avoided. If the sample cannot oe filtered in the
laboratory within 48 hours of its arrival, suffi-
cient amounts (1 ml/1 of sample) of a 2.71% solution
of mercuric chloride to give a final concentration
of 20 ppm of Hg may be added to prevent bacterial
g r ow t n.
NOTE 1: It has been reported that prevention of
bacterial growth in wacer samples can be achieved by
storing the samples in the dark.
Inter ferences
5.L Misidentification
The guidelines .set forth in this method for counting
fibrous asbestos require a positive identification
by both morphology ind crystal structure as shown by
an electron diffraction pattern. Chrysotile
asbestos has a unique tubular structure, usuallv
613
-------
showing the presence of a central canal, and
exhibits a unique characteristic electron diffrac-
tion pattern. Although halloysite fibers may show a
streaking similar to chrysotile they do not exhibit
its characteristic triple set of double spots or
5.3A layer line. It is highly improbable that a.
non-asbestiform fiber would exhibit the distin-
guishing chrysotile features. Although amphibole
fibers exhibit characteristic morphology and
electron diffraction patterns, they do not have the
unique properties exhibited by chrysotile. It is
therefore possible though not probable for misiden-
tification to take place. Hornblende is an amphi-
bole and, in a fibrous form, will be mistakenly
identified as amphibole asbestos.
It is important to recognize that a significant
variable fraction of both chrysotile and amphibole
asbestos fibers do not exhibit the required confir-
matory electron diffraction pattern. This absence
of diffraction is attributable to unfavorable fiber
orientation and fiber sizes. The results reported
will therefore be low as, compared to the absolute
number of asbestos fibers that are present.
Obscurat ion
there are large amounts of organic or amorphous
reported asbestos conten
in low
5.3 Contamination
Although contamination is not strictly considered ar.
interference, it is an important source of erroneous
results, particularly for chrysotile. The possi-
bility of contamination snould therefore always be a
cons iderat ion.
5.4 Freezing
The effect of freezing on asbestos fibers is not
knov;n cut therrj is roason to oucpect tnat fib^r
breai; down could occur an«3 roault in \ nivjlvsr fiooi
count than was orient in the original sample.
Therefore the jamplo should be transported to the
lacoratory undor conditions that would avoid
f roezi P.'} .
614
-------
6. Equipment and Apparatus
6.1 Specimen Preparation Laboratory
The ubiquitous nature of asbestos, especially chry-
sotile, demands that all sample preparation steps be
carried out to prevent the contamination of the
sample by air-borne or other source of asbestos.
The prime requirement of the sample preparation
laboratory is that it be sufficiently free from
asbestos contamination that a specimen blank deter-
mination using 200 ml of asbestos-free water yields
no more than 2 fibers in twenty grid squares of a
conventional 200 mesh electron microscope grid.
In order to achieve this low level of contamination,
the sample preparation area should be a separate
conventional clean room facility. The room should
be operated under positive pressure and have incor-
porated electrostatic precipitators in the air
supply to the room, or alternatively absolute (KZPA)
filters. There should be no asbescos floor or
ceiling tiles, transite ^heat-resistant boards, nor
asbestos insulation. V.'ork surfaces shoulj - .• 3t;i.~-
less steel or Formica or equivalent. A laminar flow
hood should be provided for sample manipulation.
Disposable plastic lab coats and disposable over-
shoes are recommended. Alternatively new shoes for
all operators should be provided and retained for
clean room use only. A mat (TacKy Mat, Liberty
Industries, 539 Deming Rd., Berlin, Connecticut
06037, or equivalent) should be placed inside the
entrance to the room to trap any gross contamination
inadvertently brought into the room from contami-
nated shoes. Normal electrical and water services,
including a distilled water supply should be
provided. In addition a source of ultra-pure water
from a still or filtration-ion exchange syscem is
desirable.
6.2 Instrumentation
6. 2. I Transr.ii j3 io.' -l-ju.-r. :i.-roocop«a . A cr ad-
mission electron microscope th^t ooeraces at
a minimum or iiO ;
-------
meter scale, concentric circles of known
radii, or other devices to measure the.
length and width of the fiber. Most modern
transmission microscopes meet the require-
ments for magnification and resolution.
An energy-dispersive X-ray spectrometer is
useful Cor the identification of-suspected
asbestiform minerals; this accessory to the
microscope, however, is not mandatory.
6.2.2 Data Processor. The large number of repeti-
tive calculations ma
-------
Screen Support
with Grid
Petri Dish
Layer of Filter Pacers
A.
Nucleoore Filter
Carbon.
;
y"^ -* >i l-i. /-* _-i
Chloroforn
r^"
r^ivar
Grid
Grid
B.
i. .".odi'Jied JaCro v;ich :*.3tl:c
A. V.'ashiny Apr.iratus
D. V.'nshiruj Process
617
-------
6.3.5 Membrane Filters.
47-ratn diameter Millipore membrane filter,
type HA, 0.45-um pore size. Used as a
Nuclepore filter support on top of glass
frit.
47-mm diameter Nuclepore membrane filter;
0.1-pm pore size. (Nuclepore Cor?., 7035
Commerce Circle, Pleasanton, CA 94566) For
filtration of water sample.
25-mm diameter Millipore membrane filter,
type HA; 0.45-um pore size. Used as
Nuclepore filter support on top of glass
frit.
25-mm diameter Nuclepore membrane filter;
0.1-um pore size. To filter dispersed ashed
Nuclepore filter.
6.3.6 Glass Vials. 30-mra diameter x SO-mm long.
For holding filter during ashing.
6.3.7 Glass Slides. 5.1-crn x 7.5-cm. For support
of Nuclepore filter during carbon
evaporation.
5.3.3 Scalpels. With disposable blades and
scissors.
6.3.9 Tweezers. Several pairs for the many
handling operations.
6.3.10 "Scotch" Doublestick tape. To hold filter
section flat on glass slide while carbon
coating.
6.3.11 Disposaole Petri dishes, 50-mm diameter, for
stocing membrane filters.
6.3.12 Static Eliminator, 500 microcuries ?o-210.
(Nuclepore Cat. No. V090POL00101) or equiva-
lent. To eliminate static charges nrcrn
membrane filters.
6.3.13 Carbon rcJ.:, spectrochemical Ly puro, 1.8"
dia., 3.6 r?.n x i.O mm neck. For carbon
coating.
6.3.14 Carbon rod sharpener. (C.it. Mo. 1204.
c^'nost F. Full am, Inc., ?. 0. oox 444,
Schenectady, NY 12301) For sharpening
carbon rods to a neck of specified length
and diameter.
3 618
-------
6.3.15 Ultrasonic Bath. (50 watts, 55 HKz). For
dispersing ashed sample and for general
cleaning.
6.3.16 Graduated Cylinder, 500 ml.
6.3.17 Spot plate.
6.3.18 10-yl Microsyringe. For administering drop
of solvent to filter section during sample
preparation.
6.3.19 Carbon grating replica, 2160 lines/mm. For
calibration of EM magnification.
6.3.20 Filter paper. S & S #539 Black Ribbon (9-cm
circles) or equivalent absorbent filter
paper. For preparing Jaffe Wick Washer.
6.3.21 Screen supports (copper or stainless steel)
12 nm x 12 mn, 200 mesh. To support
ij in JafCe Wick Washer.
6.3.22 Chlorqcorm, spectro grade, doubly
distilled. For dissolving Nuclepore filters.
6.3.23 Asbestos. Chrysotile (Canadian),
Crocidolite, Ajnosite. UICC (Union
Internationale Contre le Cancer) Standards.
Available from Duke Standards Company, 445
Sherman Avenue, Palo Alto, CA 94305.
6.3.24 Petri dish, glass {100 mm diameter x 15 r.m
high). For modified Jaffe WICK Washer.
6.3.25 Alconox. (Alcor.ox, Inc., New York, MY
10003) For cleaning glassware. Add 7.5 g
Alconox to a liter of distilled water.
6.3.26 Parafilm. (American Can Company, Neena'.-.,
WI) Use as protective covering for clean
glassware.
6.3.27 Pi pecs, oisposanLo, 3 mi and 50 -?,1.
•i.3.2J Di:>t i I lod or .!••> ion i _-v3
-------
to 100 ml distilled v/ater and dissolve by
shaking. Dilute to 200 ml with additional
water. Filter through 0.1-um Uuclepore
filter paper before using.
7. Preparation of Standards
Reference standard samples of asbestos that can be used
for quality control for a quantitative analytical method
are not available. It is, however, necessary for each
laboratory to prepare at least two suspensions, one of
chrysotile and another of a representative amphibole.
These suspensions can then be used for intra-laboratcry
control and furnish standard morphology photographs and
diffraction patterns.
7.1 Chrysotile Stock Solution.
Grind about 0.1 g of UICC chrysotile in an agate
mortar fcr several minutes, or until it appears to
be a powder. Weigh out 10 mg and transfer to a
clean 1 liter volumetric flas<, add several hundred
ml of filtered distilled water containing one ml of
a stock mercuric chloride solution and then make up
to 1 liter with filtered distilled water. To pre-
pare a working solution, transfer 10 ml of the above
suspension to another 1-liter flask, add 1 ml of a
stock mercuric chloride solution and make up tc 1
liter with filtered distilled water. This suspen-
sion contains 100 ug per liter. Finally transfer 1
ml of this suspension to a 1-liter flask, add 1 ml
of a stock mercuric cnloride solution and make up to
volume with filtered distilled water. The final
suspension will contain 5-10 MFL and is suitable for
laboratory testing.
7.2 Amphibole Stock Dispersion.
Prepare amphibole suspensions from UICC amphiccle
samples as in Section 7.1.
7.3 I j i T, zification Star. •"5i""Ji.
?r-2par-? sl^ctron microscop; : .;r:ds ror. -a in: no ~he
UICC asbestos fibors acc^': _; LT.Q co 3, rrocodur •?, ar.,i
obtain ioprocontativo phonographs ot oach rioec ty^:
ana its diffraction p.ictorr. for futuro r •? i^n ^"c "•.
3. Procedure
3.L Fi11 r11ion.
The ^oparation of the insoluble material, including
ascostiform minerals, tlirouoh filtration and suc^o-
10
620
-------
quent deposition on a membrane filter is a very
critical step in the procedure. The objective of
the filtration is not only to separate, but also to
distribute uniformly the particulate matter such
that discreet particles are deposited with a minimum
of overlap.
The volume filtered will range from 50-500 ml. In
an unknown sample the volume can not be specified in
advance because of the presence of variable amounts
of particulate matter. In general, sufficient
sample is filtered such that a very faint stain can
be observed on the filter medium. The maximum
loading that can be tolerated is 20 ug/cm2, or about
200 ug on a 47-mm diameter filter; 5 ug/cm2 is near
optimum. If the total solids content is known, an
estimate of the maximum volume tolerable can be
obtained. In a sample of high solids content, where
less than 50 ml is required, the sample should be
diluted with filtered distilled water so that a
minimum total of 50 ml of water is filtered. This
step is necessary to allow the .insoluble material to
deposit uniformly on the,filter. The filtration
funnel assembly must be scrupulously clean and
cleaned before each filtration. The filtration
should be carried out in a laminar flow hood.
NOTE 1: The following cleaning procedure has been
found to be satisfactory:
Wash each piece of glassware three times with
distilled water. Following manufacturer's recommen-
dations use the ultrasonic .bath with an Alcono:;-
water solution to clean all glassware. After the
ultrasonic cleaning rinse each piece of glassware
three times with distilled water. Then rinse eacn
piece three times with deionized water which has
been filtered through 0.1-um Nuclepore filter. Dry
in an asbestos-free oven. After the glassware is
dry, seal openings with parafilm.
8.1.1 Filtration
a. Assemble the vacuum filtration apparatus
incorporating the . I-M?, Mucleporo 'oa.:'<
-------
the container and add the entire volume to
the 47-mm diameter funnel. Apply vacuum
sufficient for filtration but gentle enough
to avoid the formation of a vortex. If a
completely unknown sample is being analyzed,
a slightly modified procedure must be
followed. Pour 500 ml of a well-mixed
sample into a 500-ml graduated cylinder and
immediately transfer the entire contents to
the prepared vacuum filtration apparatus.
Apply vacuum gently and continue suction
until all of the water has passed through
the filter. If the resulting filter appears
obviously coated or discolored, it is
recommended that another filter be prepared
in the same manner, but this time using only
200 or 100 ml of sample.
NOTE 1: Do not add more water after filtra-
tion has started and do not rinse the sides
of the funnel.
NOTE 2: Muclepore filter is basically a
hydrophobic material. The manufacturer
applies a detergent to the surface of the
filter in order to render it hydrophilic;
this process, however, does not appear to be
entirely satisfactory in some batches. Pre-
treatment of the filter in a low temperature
asher at 10 watts for 10 seconds can be used
to render the surface of the filter hydro-
philic. This process will significantly
decrease the islands of sparse deposit fre-
quently observed.
d. Disassemble the funnel, remove the
filter and dry in a covered petri dish.
8.2 Preparation of Electron Microscope Grids.
The preparation of the grid for examination in the
microscope is a critical step in the analytical
procedure. The objective is to remove the organic
filter material from the asbestos fibers with a
minimum lo.ss and movement and with a minimum broa1:-
the grid support film.
IU th-? sample contains organic matter in such
amounts that interfere with fiber counting
identification a preliminary ashing step i
r oqu i r -»d . Soe 3.5.
622
-------
8.3 Nuclepore Filter, Modified Jaffe Wick Technique.
8.3.1 Preparation of Modified Jaffe Washer
Place three glass microscope slides (75 mm x
22 mm) one on top of the other in a petri
dish (100 mm x 15 mm) along a diameter.
Place 14 S & S #589 Black Ribbon filter
papers (9-cm circles) in the petri dish over
the stack of microscope slides. Place three
copper mesh screen supports (12 mm x 12 mm)
along the ridge formed by the stack of
slides underneath the layer of filter
papers. Place an EM specimen grid on each
of the screen supports. See Fig. 1.
NOTE 1: A stack of 30-40 S & S filters
(7-cm circle) can be substituted for the 14
filters and microscope slides in preparing
the Jaffe washer.
8.3.2 Vacuum Filtration Unit
Assemble the vacuum filtration unit. Placa
a 0.45-ura Millipors filter type KA on the
glass frit and then position a 0.1-um
Nuclepore filter, shiny side up, on top of
the Millipore filter. Apply suction to
center the filters flat on the frit. Attach
the filter funnel-and shut off the suction.
8.3.3 Sample Filtration
See 3.1.1.
8.3.4 Sample Drying
Remove the filter funnel and place the
Nuclepore filter in a loosely covered petri
dish to dry. The petri dish containing the
filter may be placed in an asbestos-free
oven at 45° C for 30 minutes to shorten
the drying time.
3.3.5 Selection of Section for Carbon Coating
Using a small pair of scissors or sharp
scalpel cut out a retangular section oc the
Nuclepore filter. The minimum approximate
dimensions snould be 15 n:m long and 3 :nm
•.. ijj. ."xw..., ... ^ec _ .jr. r.^ar tne perimeter of
the filtration area.
13
623
-------
8.3.6 Carbon Coating the Filter
Tape the two ends of the selected filter
section to a glass slide using "Scotch"
tape. Take care not to stretch the filter
section. Identify the filter section using
a china marker on the slide. Place the
glass slide with the filter section into the
vacuum evaporator. Insert the necked carbon
rod and, following manufacturer's instruc-
tions, obtain high vacuum. Evaporate the
neck, with the filter section rotating, at a
distance of approximately 7.5 cm from the
filter section to obtain a 30-50 nm layer of
carbon on the filter paper. Evaporate the
carbon in several short bursts rather than
continuously to prevent overheating the sur-
face of the Nuclepore filter.
NOTE 1: Overheating the surface tends to
crosslink the plastic, rendering the filter
dissolution in chloroform difficult.
NOTE 2: The thickness of the carbon film
can be monitored by placing a drop of oil on
a porcelain chip that is placed at the same
distance from the carbon electrodes as the
specimen. Carbon is not visible in the
region of the oil drop thereby enabling the
visual estimate of the deposit thickness by
the contrast differential.
8.3.7 Grid Transfer
Remove the filter from the vacuum evaporator
and cut out three sections somewhat less
than 3 mm x 3 mm and such that the square of
Nuclepore fits within the circumference o£
the grid. Pass each of the filter sections
over a static eliminator and then place each
of the three sections carbon-sice down on
separate specimen grids previously placed in
the modified Jaffe Washer. Using a micro-
syringe, place a 10-ijl drop of chloroform on
each filter section resting on a grid and
then saturate the filter pad until pooling
of the solvent occurs below the ridge forme-:
by the glass .slides inserted under th* L.?.y:>r
of filter papers. Place the cover on th-j
oetri dish and allow the grids to remain i;i
the washer for approximately 24 hour - .
not allow the chloroform to completely
evaporate before the grids are removed. To
remove the grids from the washer lift the
14
624
-------
screen support with the grid resting upon it
and set this in a spot plate depression to
allow evaporation of any solvent adhering to
the grid. The grid is now ready for
analysis or storage.
8.4 Electron Microscopic Examination
8.4.1 Microscope Alignment and Magnification
Calibration
Following the manufacturer's recommendations
carry out the necessary alignment procedures
for optimum specimen examination in the
electron microscope. Calibrate the
routinely used magnifications using a carbon
grating replica.
NOTE 1: Screen magnification is not
necessarily equivalent to plate
magni f icat ion .
3.4.2 Grid Preparation Acceptability
After inserting the specimen into the micro-
scope adjust the magnification low enough
(300X-1000X) to permit viewing complete grid
squares. Inspect at least 10 grid squares
for fiber loading and distribution, debris
contamination, and carbon film continuity.
Reject the grid for counting if:
1) The grid is too 'heavily loaded with
fibers to perform accurate counting and
diffraction operations. A new sample? prepa-
ration either from a smaller volume of water
or from a dilution with filtered distilled
water must then be prepared.
2) The fiber distribution is noticeably
uneven. A new sample preparation is
required .
3) The debris contamination is too
to perform accurate counting .and diffraction
operations. If the do brio i 3 Largely
organic the filtor must be ashed and ro/ji^-
por3i?t.l (.»oe 3.5) . It: inorganic the ^a.npLo
must be diluted ,ind .again piepared.
• ' Tho majority of gi'iJ squares examined
have broken carbon films. A different grid
1
625
-------
preparation from the same initial filtration
must be substituted.
3.4.3 Procedure for Fiber Counting
There are two methods commonly used for
fiber counting. In one method (A) 100
fibers, contained in randomly selected
fields of view, are counted. The number of
fields plus the area of a field of view must
be known when using this method. In the
other method (B), all fibers (at least 100)
in several grid squares or 20 grid squares
are counted. The number of grid squares
counted and the average area of one grid
square must be known when using this method.
NOTE 1: The method to use is dependent upon
the fiber loading on the grid and it is left
to the judgement of the analyst to select
the optimum method. The following guide-
lines can be used: If it is estimated that
a grid square (30 urn x 80 urn) contains
50-100 fibers at va screen magnification of
20000X it is convenient to use the field-of-
view counting method. If the estimate is
less than 50, the grid square method of
counting should be chosen. On the other
hand, if the fiber count is estimated to be
over 300 fibers per grid square, a new grid
containing less fibers must be prepared
(through dilution or filtration of a smaller
volume of water).
3.4.3A Field-of-View Method
After determining that a fiber count can be
obtained using this method adjust the screen
magnification to 15,-20,OOOX. Select a
number of grid squares that would be as
representative as possible of the entire
analyzable grid surface. From each of thes.j
squares select a sufficient number of field.*
of view for floor co.:~. ': in ~:. The number of
fields of view per <]•: id square i3 dependent
upon the fiber Loading. If more than one
field of view per grid squar3 is selected,
jean the grid opening orthogonally in an
arbitrary pattern which prevents overlapping
of t'ielJj ot" view. Carry out the analysis
by counting, measuring and identifying (3ee
8.4.4) approximately 50 fibers on each of
two gr ids.
1*5
626
-------
The following rules should be followed when
using the field of view method of fiber
counting. Although these rules were derived
for a circular field of view they can be
modified to apply to square or rectangular
designs.
1) Count all fibers contained within the
counting area and not touching the circum-
ference of the circle.
2) Designate the upper right-hand quadrant
as I ana number in clockwise order. Count
all fibers touching or intersecting the arc
of quadrants I or IV. Do not count fibers
touching or intersecting the arc of quad-
rants II or III.
3) If a fiber intersects the arc of both
quadrants III and IV or I and II count it
only if the greater length was outside the
arc of quadrants IV and I, respectively.
4) Count fibers intersecting the arc of
both quadrants I and III but not those
intersecting the arc of both II and IV.
These rules are illustrated in Fig. 2.
8.4.3B Grid Square Method
After determining that a fiber count can be
obtained using this, method adjust the screen
magnification to 15,-20,OOOX. Position the
grid square so that scanning can be started
at the left upper corner of the grid
square. While carefully examining the grid,
scan left to right, parallel to the upper
grid bar. When the perimeter of the grid
square is reached adjust the field of view
down one field width and scan in the oppo-
site direction. The tilting section of the
fluorescent screen may be used conveniently
as the field of view. Examine the square
until all the area has bo-?n covered. The
analysis should no carried out by count inr;,
measuring and identify ing (soe 3.4.4)
approximately 50 Libers on t?ach of t-.;o grid.;
or until 10 griJ squares on each oC two
qri'Ja have been counted. Do not count
fibers intersecting a gric ^r.
17
627
-------
IV
III
II
Counted
Mot Counted
•j 2. Illustration of Cour.tir.y Rules
for Fio Ul-of-V low Method
i:
628
-------
8.4.4 Measurement and Identification
Measure and record the length and width of
each fiber having an aspect ratio greater
than or equal to three. Disregard obvious
biological, bacteriological fibers and
diatom fragments. Examine the morphology of
each fibber using optical viewing if
necessary. Tentatively identify, by refer-
ence to the UICC standards, chrysotile or
possible amphibole asbestos. Attempt to
obtain a diffraction pattern of each fiber
utilizing the shortest camera length
possible. Move the suspected fiber image to
the center of the screen and insert a suit-
able selected area aperture into the
electron beam so that the fiber image, or a
portion of it, is in the illuminated area.
The size of the aperture and the portion of
the fiber should be such that particles
other than the one to be examined are
excluded from the selected area. Observe
the diffraction ^pattern with the 10X binocu-
lars. If an incomplete diffraction pattern
is obtained move the particle image around
in the selected area to get a clearer
diffraction pattern or to eliminate possible
interferences from neighboring particles.
Determine whether, or not the fiber is chry-
sotile or an amphibole by comparing the
diffraction pattern obtained to the diffrac-
tion patterns of known standard asbestos
fibers. Confirm the tentative identifi-
cation of chrysotile and amphibole asbestos
from their electron diffraction patterns.
Classify each fiber as chrysotile, amphi-
bole, non-asbestos, no diffraction or
ambiguous.
MOTE 1: It is convenient to use a tape
recorder during the examination of the
fibers to record all pertinent data. This
information can then bo summarized on data
sheets oc punched cards for subsequent auto-
matic data processing.
:JOTZ 2: Chrysotile fibers occur as single
fibrils, or in bundles. The fibrils gene-
rally show a tubular structure with a hollow
canal, altnough the absence of the canal
does not rule out its identification.
Amphibole .isbostos fibers usually exhibit a
lath-like structure with irregular ends, but
L'J
629
-------
occasionally will resemble chrysotile in
appearance.
NOTE 3: The positive identification of
asbestos by electron diffraction requires
some judgment on the part of the analyst
because some fibers give only partial
patterns. Chrysotile shows unique prominent
streaks on the layer lines nearest the
central one and a triple set of double spots
on the second layer line. The streaks and
the set of double spots are the distin-
guishing characteristics of chrysotile
required for identification. Amphibole
asbestos requires a more complete diffrac-
tion pattern to be positively identified.
As a qualititative guideline, layer lines
for amphibole, without the unique streaks
(some streaking may be present) of chryso-
tile, should be present and the arrangement
of diffraction spots along the layer lines
should be consistent with the amphibole
pattern. The pattern should be distinct
enough to establish these criteria.
NOTE 4: Chrysotile and thin amphibole
fibers may undergo degradation in an elec-
tron beam; this is particularly noticeable
in small fibers. It may exhibit a pattern
for a 1-2 seconds and disappear and the
analyst must be alert to note the charactar-
istic features.
NOTE 5: An ambiguous fiber is a fiber chat
gives a partial electron diffraction patter-
resembling asbestos, but insufficient to
provide positive identification.
8.4.5 Determination of Grid Square Area
Measure the dimensions of several represen-
tative grid squares from each batch of griis
with an optical microscope. Calculate the
average area of a grid square. This should
be done to compensate for variability in
grid square dimensions.
3.5 Ajh i n
Some samples contain aut f ic iontly high Level:; of
organic material that an ashing step is required
beL'ore fiber identification and counting can be
car r ied out .
23
630
-------
Place the dried Nuclepore filter paper containing
the collected sediment into a glass vial (23 mm dia-
meter x 80 mm high). Position the filter such that
the filtration side touches the glass wall. Place
the vial in an upright position in the low tempera-
ture asher. Operate the asher at 50 watts (13.55
MHz) power and 2 psi oxygen pressure. Ash the
filter until a thin film of white ash remains. The
time required is generally 6 to 8 hours. Allow the
ashing chamber to slowly reach atmospheric pressure
and remove the vial. Add 10 ml of filtered
distilled water to the vial. Place the vial in an
ultrasonic bath for 1/2 hour to disperse the ash,
Dilute the sample if required.
Assemble the 25-mm diameter filtering apparatus.
Center a 25-mm diameter .1-ym Nuclepore filter (with
the 0.45-um Millipore backing) on the glass frit.
Apply suction and recenter the filter if necessary.
Attach the filter funnel and turn off the suction.
Add the water containing the dispersed ash from the
vial to the filter funnel. Apply suction and filter
the sample. After drying this filter it is ready to
be used in preparing sample grids as in 8.3.
NOTE 1: In specifying a 25-mm diameter filter it is
assumed that the ashing step is necessary mainly
because of the presence of organic material and that
the smaller filtering area is desirable from the
point of view of concentrating the fibers. If the
sample contains mostly inorganic debris such that
the smaller filtering area will result in over-
loading the filter, the 47-mm diameter filter should
be used.
NOTE 2: It will be noted that a 10-ml volume is
filtered in this case instead of the minimum 50~ml
volume specified in 8.1. These volumes are consis-
tent when it is considered that there is approxi-
mately a 5-fold difference in effective filtration
area between the 25-mm diameter and 47-mm diameter
filters.
3.o Determination of Blank Level
Carry out a blank J»? term inat ton with o.ich batch or
samplos prepared, but a minimum of one per week.
Filter 3 -Jrosh .supply (500 ml) of distilled,
de ionized wator through a clean 0.1-'..m membrane
filter. Filter 200 ml of this w.iter through a
O.L-um Nuclepore filter, prepare the electron micro-
scope grid, and count exactly as in the procedures
8.1 - 3.4. Examine 20 grid squares and record this
number of fibers. A maximum of two fibers in 20
grid squares 13 acceptable for the blank sample.
631
-------
NOTE 1: The monitoring of the background level of
asbestos is an integral part of the procedure. Upon
initiating asbestos analytical work, blank samples
must be run to establish the initial suitability of
the laboratory environment, cleaning procedures, and
reagents for carrying out asbestos analyses.
Analytical determinations of asbestos can be carried
out only after an acceptably low level of contami-
nation has been established.
9. Calculations
9.1 Fiber Concentrations
Grid Square Counting Method - If the Grid Square
Method of counting is employed, use the following
formula to calculate the total asbestos fiber
concentration in MFL.
C = (F x Af)/(Ag x V0 x 1000)
If ashing is involved use the same formula but
substituting the effective filtration area of the
25-mm diameter filter for Af instead of that for
the 47-mm diameter filter. If one-half the filter
is ashed, multiple C by tv/o.
C = Fiber concentration (MFL)
F = average number of fibers per grid opening
Ac = Effective filtration area of filter
paper (mnr) used ia grid preparation for
fiber counting
Aq = Average area of one grid square {mnr)
VQ = Original volume of sample filtered (mi)
Field-o£-View Count ina Method - If the Field-of-View
Method of counting is employed use the following
formula to calculate the total asbestos fiber
concentrations (MFL)
C - (F x Af x 1000) '(Av :< Vo)
[£ Jo.ninq i 3 involved use the same formula but
su'b.31 i tut inq tho otroctivo ciltration ,ir
-------
F = Average number of fibers per field of
view
Af = Effective filtration area of filter
paper (ram2) used in grid preparation for
fiber counting
Av = Area of one field of view ( um 2)
VQ = Original volume of sample filtered (ml)
9.2 Estimated Mass Concentration
Calculate the mass (ug) of each fiber counted using
the following formula:
M = L x W2 x D x 10-6
If the fiber content is predominantly chrysotile,
the following formula may be used:
M = 1 x L x W2 x. D x 10-°
4
where M = mass (ug)
L = length (ym)
W = width (urn)
D = density of fibers fc/cm3)
Then calculate the mass concentration (ug/D
employing the following formula.
l\, = C x Mf x 10
where M = mass concentration (yg/1)
C = fiber concentration (MFL)
FL s mean mass per fiber (ug)
To calculate H- use the following formula:
whcro M/ =« mass of each fiber, respectively
n = number of fibers counted
633
-------
NOTE 1: Because many of the amphibole fibers are
lath shaped rather than square in cross section the
computed mass will tend to be high since laths will
in general tend to lie flat rather than on edge.
NOTE 2: Assume the following densities: Chrysotile
2.5, Amphibole 3.25
9.3 Aspect Ratio
The aspect ratio for each fiber is calculated by
dividing the length by the width.
10. Reporting
10.1 Report the following concentration as MFL
a. Total fibers
b. Chrysotile
c. Arophibole
10.2 Use two significant figures for concentrations
greater than 1 MFL, and one significant figure for
concentrations less than 1 MFL.
10.3 Tabulate the size distribution, length and width.
10.4 Tabulate the aspect ratio.distribution.
10.5 Report the calculated mass as ug/1.
10.6 Indicate the detection limit in MFL.
10.7 Indicate if less than five fibers were counted.
10.8 Include remarks concerning pertinent observations,
(clumping, amount of organic matter, debris) amount
of suspected though not identifiable as asbestos
fibers (ambiguous).
11. Precision
11.1 Intra-Laboratory
The precision that is obtained within an individual
laootratory is dependent upon the number of fibers
counted. It" LOO fibers are counted and the loading
is at least 3.5 fibers/grid square, computer
modeling of the counting procedure shows a relative
standard deviation of abouc LOs can be expected.
21
634
-------
In actual practice some degradation from this
precision will be observed but should not exceed +
15% if several grids are prepared from the same
filtered sample. The relative standard deviation of
analyses of the same water sample in the same labo-
ratory will increase due to sample preparation
errors and a relative standard deviation of about
about + 25 - 35% will occur. As the number of
fibers counted decreases, the precision will also
decrease approximately proportional to /N where N is
the number of fibers counted.
Based upon the analysis of one laboratory utilizing
a different analyst for each of three water samples,
intra-laboratory precision data is presented in
Table 1.
Table 1. Intra-Laboratory Precision
Sample Number of Mean Fiber Precision,
Type Sample Concentration Relative
Aliquots MFL (millions of Standard
Analyzed asbestos fibers/1) Deviation
Chrysotile 25 23 37%
(UICC)
Crocidolite 20 8 36%
(UICC)
Taconite 20 16 24%
(raw water)
11.2 Inter-Laboratory
Based upon the analysis by various government and
private industrial laboratories of filters prepared
from nine v/ater samples, inter-laboratory precision
data of the method is oresented in Table 2.
635
-------
Table 2. Inter-Laboratory Precision
Sample
Type
Chrvsotile
Amphibole
Niunber of
Labs
Reporting
10
9
11
9
9
3
11
4
14
Mean Fiber
Concentration
MFL (millions of
asbestos fibers/1)
877
119
59
31
23
25
139
95
36
Precision,
Relative
Standard
Deviation
35%
43%
41%
65%
32%
35%
50%
52%
66%
12. Accuracy
12.1 Fiber Concentrations
As no standard reference materials are available,
only approximate estimates of the accuracy of the
procedure can be made. At 1 MFL, it is estimated
that the results should be within a factor of 10 of
the actual asbestos fiber content.
This method requires the positive identification of
fiber to be asbescos as a means for its quantitative
precludes t
determination. As the state-of-the-art
positive identification of -all of the asbestos fibers
present, the results of this method, as expressed as
MFL, will be biased on the low side and assuming no
fiber loss represent 0.4 - 0.3 of the total asbestos
fibers present.
12.2 Mass Concentrations
no stan-
As in the case of the fiber concentrations,
dard samples of the size distribution found
are available. The accuracy of the mass
nation should be somewhat better than
determination
of the larger
portion of the mass, are
counted. This will reduce the bias of low
duo to d; f L icuLc i-.?'5 in identification. At
the assumption that the thickness of
the width will result in a positive error
determining tho volume ot the fiber and thus give
high results tor the mass.
time,
determi-
the fiber
if a statistically significant number
fibers, which contribute the major
identified, measurod, .ind
results
the same
the fibor
in
636
-------
13. Suggested Statistical Evaluation of Grid Fiber Counts
13.1 Since the fiber distribution on the sample filter,
resulting from the method of filtration, has not been
fully characterized, the fiber distribution obtained
on the electron microscope grids for each sample
should be tested statistically against an assumed
distribution and a measure of the precision of the
analysis should be provided.
13.2 Assume that the fibers are uniformly and randomly
distributed on the sample filter and grids. One
method for confirming this assumption is given below.
13.3 Using the chi-square test, determine whether the
total number of fibers found in individual grid
openings are randomly and uniformly distributed among
the openings, by the following formula:
N
X2 =
13.4
where X: - chi-square statistic
N = number of grid openings examined for the
sample
n.- = total number of fibers found in each
respective grid opening
n = total number of fibers found in II grid
openings
p£ - ratio of the area of each respective
grid opening to the sun of the areas of all
grid openings examined
NOTE 1: If an average area for the grid squares has
been measured as outlined in 3.4.5, the term np•
represents the mean fiber count per grid square.'
If the value for X: exceeds the value listed in
statistical tables for the 0.1$ significance level
with M-l degrees of freedom, tne fibers are not
considered to be uniformly and randomly distribute-
among the grid openings. IP. this case, it is
advisable to try to improve the uniformity of fib-v:
deposition by filtering another aliquot of the sample
and repeating tu: .analyr, is.
If uniformity and randomness of fiber deposition on
the microscope grids has oeen demonstrated as in
13.3, the 95'i confidence interval about the mean
637
-------
fiber counts for chrysotile, amphibole, and total
asbestos fibers may be determined using the following
formulae:
(1) Sc =
N N
V x2"- f V y ^
Z. xv \ 2 x-)
N(N-l)
2
1/2
where Sc = standard deviation of the chrysotile
fiber count
N = number of grid openings examined for the
sample
Xv = number of chrysotile fibers in each grid
opening, respectively
Obtain the standard deviations of the fiber counts
for anphibole asbestos fibers and for total asbestos
fibers by substituting the corresponding value of X
into equation ( 1) .
ts
(2) X = X + -~
- tSc
(3) X. = X - -=°-
L y'
where X = upper value of 95% confidence interval
for chrysotile
XL = lower value of 95% confidence interval
for chrysotile
X = average number of fibers per grid opening
t =» value listed in t-distr ibut ion tables at
the 95% confidence l^vel for a two tailed
distribution with N-L degree of freedom
S^ = standard deviation of the fioer count::
for chrysotilo
a number of grid openings examined for the
sample
The values of Xu and XL can be converted to concen-
trations in millions of fibers per liter using the
formula in sec_tion 9 and substituting either Xy or XL
for the term F.
638
-------
Obtain the upper and lower values of the 95% confi-
dence interval for amphibole asbestos fibers and
total asbestos fibers by substituting the corres-
ponding values of X and S into equations (2) and (3).
Report the precision of the analysis, in terms of the
upper and lower limits of the 95% confidence
interval, for chrysotile, amphibole, and total
asbestos fiber content. If a lower limit is found to
be negative, report the value of the limit as zero.
SELECTED BIBLIOGRAPHY
Bearaan, D. R. and D. M. File.
Asbestos Fiber Concentrations.
Quantitative
Anal. Chem.
Determination of
48(1): 101-110, 1976
Lishka, R. J., J. R. Millette, and E. F. McFarren. Asbestos
Analysis by Electron Microscope. Proc. AWWA Water Quality Tech.
Conf. American Water Works Assoc., Denver, Colorado XIV-1 -
XIV-12, 1975.
Mi'llette, J. R. and E. F. McFarren. EDS of Waterborne Asbestos
Fibers in TEM, SEM and STEM. Scanning Electron Microscopv/1976
(Part III) 451-460, 1976.
Cook, P. M., I. B. Rubin, C. J. Maggiore, and W. J. Nicholson.
X-ray Diffraction and Electron Beam Analysis of Asbestiform
Minerals in Lake Superior Waters. Proc. Inter. Conf. on
Environ. Sensing and Assessment 34(2): 1-9, 1976.
McCrone, W.
10-13, 1974
C. and I. M. Stewart. Asbestos. Amer. Lab.
(4;
Mueller, P. K., A. E.
Asbestos Fiber Atlas.
Alcocer, R. L. Stanley, and G. R. Smith.
U.S. Environmental Protection Agency
Technology Series, EPA 650/2-75-036, 1975.
Glass, R. W. Improved Methodology for Determination of Asbestos
as a Water Pollutant. Ontario Research Foundation Report, April
30, 1976. Mississauga, Ontario, Canada.
Samudra, A. V. Optimum Procedure for Asbestos Fibers Identifi-
cation from Selection Area Electron Diffraction Patterns in a
Modern Analytical Electron Microscope Using Tilted Specimer.3.
Scanning Electron Microscopy, Vol. I, Proceedings of t;he Work-
shop on Analytical Electron Microocopy, March, 1977. Chicago,
111inoi 3.
Chatfield, E. J., R. W. Gl,-. ;, :.- j .",. J. Dillon. Preparation ot
Water Samples tor Asbestos Fiber Counting by Electron Micros-
copy. EPA Report EPA-600/4-73-011, January 1978.
639
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Asher, I. M. and. P. P. McGrath. Symposium on Electron Micros-
copy of Microfibers. Proceedings of the First FDA Office of
Science Summer Symposium, August 1976. Pennsylvania State
University.
Chopra, K. S. "Interlaboratory Measurements of Amphibole and
Chrysotile Fiber Concentrations in Water." Journal of Testing
and Evaluation, JTEVA, Vol. 6, No. 4, July 1978, pp. 241-247.
National Bureau of Standards Special Publication 506.
Proceedings of the Workshop on Asbestos: Definitions and
Measurement Methods held at NBS, Gaithersburg, MD, July 18-20,
1977. (Issued March 1978).
*U.S. GOVERNMEHT PRINTING OFFICE: 1982-0-361-085/^58
640
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