EPA-650/2-74-085-b



SEPTEMBER 1974
                                Environmental Protection  Technology  Series



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                                      EPA-650/2-74-085-b
CONTROL OF SULFUR  DIOXIDE EMISSIONS
  FROM COPPER SMELTERS: VOLUME  II -
     HYDROGEN  SULFIDE PRODUCTION
      FROM COPPER  CONCENTRATES
                        by

               C.A. Rohrmann andH.T. Fullam

             Battelle Pacific Northwest Laboratories
                   Battelle Boulevard,
                Richland, Washington 99352

                  Contract No. 68-02-0025
                Program Element No. 1AB013
                  ROAPNo. 21ADC-056

               EPA Project Officer: L. Stankus

                Control Systems Laboratory
             National Environmental Research Center
           Research Triangle Park, North Carolina 27711

                     Prepared for

            OFFICE OF RESEARCH AND DEVELOPMENT
           U.S. ENVIRONMENTAL PROTECTION AGENCY
                WASHINGTON, D.C.  20460

                     September 1974

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This report has been reviewed by the Environmental Protection Agency and
approved for publication.  Approval does not signify that the contents
necessarily reflect the views and policies of the Agency, nor does mention
of trade names or commercial products constitute endorsement or recommen-
dation for use.
                                11

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                                   -11-
                            TABLE OF CONTENTS

LIST OF FIGURES	iv
LIST OF TABLES	vi
ABSTRACT   	    1
SUMMARY	    3
INTRODUCTION	    6
BACKGROUND 	    9
PROCESS DESCRIPTION 	   14
EXPERIMENTAL PROGRAM   	   18
   COPPER CONCENTRATES STUDIED  	   18
   NEUTRAL ROASTING OF CONCENTRATES   	   21
   ACID LEACHING OF NEUTRAL ROASTED CONCENTRATES  	   22
   ANALYTICAL PROCEDURES  	   25
RESULTS AND DISCUSSION 	   28
   NEUTRAL ROASTING 	   28
      Composition of Neutral Roasted Concentrates 	   28
      X-Ray Analysis of Neutral Roasted Concentrates 	   33
      Thermal Analysis of Concentrates   	   39
      Trace Impurities	48
      Sulfur Dioxide Formation  	   49
   ACID LEACHING	51
      Concentrate Type	54
      Concentrate Composition   	   57
      Operating Variables 	   62
      Fate of Impurities During Leaching 	   70
      Process Optimization   	   76
ECONOMIC ANALYSIS OF PROCESS 	   78
   NEUTRAL ROASTING SYSTEM   	   84
   LEACHING OPERATION  	   87
   ACID RECOVERY SYSTEM	90
   S02 RECOVERY SYSTEM	93

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                                  -m-
                        TABLE  OF  CONTENTS  (contd)

   CONVERTER OPERATION 	    94
   TOTAL PROCESSING COSTS 	    97
CONCLUSIONS	102
RECOMMENDATIONS  	   106
REFERENCES	107

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                                   -IV-
                              LIST OF FIGURES

 1     Schematic Flow Diagram for Acid Leaching  Copper Concentrates.   .   15
 2     Equipment for Leaching Neutral-Roasted Concentrates
      in Hydrochloric Acid	23
 3     Typical  Sulfide Titration Curve Using Sulfide  Electrode   ...   26
 4     Mineral  Compositions in the Fe-Cu-S System   	   35
 5     Thermogravimetric Analysis of Copper Concentrates  in
      Helium	40
 6     Thermograms for Concentrates in Helium  	   41
 7     Effect of Heating Rate on the Thermal Decomposition of
      Morenci  Concentrate  	   43
 8     Effect of Temperature on Roasting of Morenci Concentrate  ...   44
 9     Differential  Thermal Analysis Curves for  Copper Concentrates
      Heated in Purified Helium	   .   46
10     Thermograms for Morenci Concentrate 	   47
11     Typical  Leaching Data for Neutral  Roasted Morenci  Concentrate   .   52
12     Reactivity of Neutral Roasted Morenci Concentrate   	   59
13     Effect of Initial Acid Concentration on Leaching of Neutral
      Roasted  Morenci Concentrate   	   64
14     Effect of Initial Acid Concentration on Dissolution of
      Neutral  Roasted Morenci Concentrate 	   65
15     Effect of Initial Acid Concentration on Leaching of
      Neutral  Roasted Pima Concentrate 	   66
16     Effect of Initial Acid Concentration on the  Leaching  of
      Neutral  Roasted Anaconda Concentrate	67
17     Effect of Excess Acid on Leaching of Neutral Roasted
      Morenci  Concentrate  	   68
18     Effect of Excess Acid on Leaching of Neutral Roasted
      Pima Concentrate	69
19     Effect of Temperature on Leaching of Neutral Roasted
      Morenci  Concentrate  	   71
20     Effect of Temperature on Leaching of Neutral Roasted  Pima
      Concentrate	72
21     Effect of Temperature on Leaching of Neutral Roasted
      Anaconda Concentrate 	   73

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                                   -V-
                          LIST OF FIGURES (contd)
22    Effect of Reaction Temperature on the Leaching of Neutral
      Roasted Morenci  Concentrate 	  74

23    Process Flow Diagram for 300 T/D Copper Smelter Using
      Morenci Concentrate Plus Pyrite Concentrate	80

24    Process Flow Diagram for a 300 T/D Copper Smelter Using
      Pima Concentrate Plus Pyrite Concentrate   	  81

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                                  -VI-
                             LIST  OF TABLES

    I    Composition of Concentrates  Evaluated 	    19
   II    Composition of Neutral  Roasted Concentrates  	    29
  III    Effect of Temperature on Composition of Neutral
        Roasted Morenci Concentrate   	    31
   IV    Composition of Neutral  Roasted Pima  and Anaconda
        Concentrates Prepared at Various  Temperatures   	    32
    V    Compounds in the Fe-Cu-S System that are Stable
        at Low Temperatures	34
   VI    The X-Ray Diffraction Patterns for Samples of  Morenci
        Concentrates which were Neutral Roasted at 800C  in
        Flowing Argon on the Thermobalance 	    37
  VII    X-Ray Diffraction Patterns for Neutral  Roasted Pima
        Concentrate Prepared Under Various Conditions   	    38
 VIII    Effect of Neutral Roasting at 800C  on  the  Impurities
        in Anaconda Concentrate	48
   IX    Material  Balance Data for  a  Typical  Leaching Experiment
        Using Neutral Roasted Morenci Concentrate	53
    X    Leaching Data for Neutral  Roasted Concentrates 	    56
   XI    Effect of Concentrate Composition on Leaching  ,	58
  XII    Effect of Composition on Leaching of Neutral Roasted
        Pima and Anaconda Concentrates	61
 XIII    Leaching of Neutral Roasted  Pyrite-Copper Concentrate
        Mixtures	61
  XIV    Fate of Impurities During  Leaching of Anaconda
        Concentrate	75
   XV    Effect HpS Treatment on Impurities in the Leach  Solution.   .    76
  XVI    Bases Used in Preparing Flow Diagrams	82
 XVII    Basis for Estimating Plant Operating Costs   	    83
XVIII    Capital Cost for Neutral Roasting Operation  	    85
  XIX    Operating Cost for Neutral Roasting  Operation   	    86
   XX    Capital Cost for the Leaching Operation	88
  XXI    Operating Costs for the Leaching Operation   	    89
 XXII    Capital Costs for Acid Recovery System	91
XXIII    Operating Costs for Acid Recovery System 	    92

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                                   -V11-
                          LIST OF TABLES  (contd)

  XXIV   Capital  Costs for S02 Recovery System 	    95
   XXV   Operating Costs for S0? Recovery System   	    96
  XXVI   Capital  Costs for Converter Operation 	    98
 XXVII   Operating Costs for Converter Operation   	    99
XXVIII   Capital  Costs for a 300 Ton/Day  Copper Smelter  	   100
  XXIX   Yearly Operating Costs for a 300 Ton/Day  Copper
         Smelter	101

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                                    -1-
         CONTROL OF SULFUR DIOXIDE EMISSIONS FROM COPPER SMELTERS:
     VOLUME  II - HYDROGEN SULFIDE PRODUCTION FROM COPPER CONCENTRATES
                                 ABSTRACT

     A laboratory investigation has been made of a modified copper smelting
process which provides a solution to the sulfur dioxide air pollution prob-
lem and produces blister copper, elemental sulfur, and iron oxide without
loss of the precious metals.  Preliminary economic evaluation of the process
appears favorable with good prospects for further improvements when com-
pared with conventional processes provided with equivalent air pollution
abatement capabilities.
     The process would involve (1) neutral roasting of pyritic copper con-
centrates to convert the contained iron into an acid-soluble form with
evolution of some elemental sulfur in this step, (2) hydrochloric acid
leaching of the roasted concentrate to dissolve the iron with simultaneous
hydrogen sulfide generation and production of an enriched copper sulfide
residue, (3) converting the copper sulfide residue to blister copper by
conventional means, (4) reducing the sulfur dioxide formed in the convert-
ing step to elemental  sulfur with hydrogen sulfide from the leaching step,
and (5) processing the iron chloride leach solution to regenerate hydro-
chloric acid and to yield a marketable iron oxide.   Step 1  (for pyrite con-
version), as well  as steps 3, 4 and 5, are regarded as present industrial
state-of-the-art.   Step 2, based on the results from the principal  focus of this
laboratory study,  would involve relatively straightforward chemical  engineering
development for scale-up to pilot-plant scale to demonstrate feasibility.
     Principal  advantages of the process include:   (1) hydrogen sulfide
production without requiring the utilization of a  specific  chemical  reduc-
tant such as costly natural gas or coke; (2) elimination of the costly
investment and  operations of reverberatory furnaces with their accompanying
high fuel  demand,  dilute sulfur dioxide flue gas releases and concurrent

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                                   -2-
formation of massive amounts  of high-temperature  solid wastes;  (3) major
reductions in the requirements  for slag-forming minerals; and  (4) probable
increased overall recovery of copper and  precious metals from  a  specific
ore body.

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        CONTROL OF SULFUR  DIOXIDE  EMISSIONS  FROM  COPPER SMELTERS:
     VOLUME II  - HYDROGEN  SULFIDE  PRODUCTION FROM COPPER CONCENTRATES
                                 SUMMARY

     As a result of the laboratory work done  under  this  project,  it  has
been determined that:   (1)  by neutral  roasting  of pyritic  copper  concen-
trates at 800C or higher a large fraction of the iron in  such  concentrates
is converted to an acid-soluble sulfide form; (2) by dissolution  of  the
solubilized iron in 4-5N. hydrochloric  acid at 80-85C large  quantities of
concentrated hydrogen sulfide can be readily  obtained;  (3) an enriched
copper sulfide still containing the precious  metals is readily  separated
from the acidic iron chloride solution; and (4) the small  amounts of copper
which are also dissolved can be readily recovered by treatment  of the
cooled acidic-iron chloride solution with hydrogen  sulfide.   This treatment
also assures the retention of the precious metals,  principally  silver, with
the copper sulfide residues.
     The regeneration and recovery of the hydrochloric acid  from the
ferrous chloride solution as produced above is regarded  as a present state-
of-the-art commercial process applied on  a substantial  and increasing scale
in the steel pickling industry.
     The conversion of the moderately-concentrated  sulfur dioxide evolved
from conventional copper smelter converters to elemental sulfur by reaction
with the hydrogen sulfide produced as indicated above is also regarded as  a
present state-of-the-art process, the so-called Claus process.   However,  an
alternative means utilizing  the  U.S. Bureau of Mines Citrate process, which
is  in  a fairly  advanced state of development, may also be employed for
elemental  sulfur production.
      The production of blister copper from the enriched copper sulfide con-
centrate obtainable by the  above treatment is  regarded as a present state-
of-the-art process  which can be  accommodated in copper smelter converters
of  the existing design.

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                                   -4-
     Thus the technical,  or at least the chemical,  feasibility  of  the  pro-
cesses necessary for the  recovery of elemental  sulfur  in  high yield  and
quality from copper smelter flue gases  without  need for a specific and
increasingly more costly  chemical reductant such  as natural  gas or coke
has been demonstrated on  the laboratory scale.  This is regarded as  the
positive answer to the principal original  objective of this  investigation.
     In reviewing the composition of representative pyritic  copper concen-
trates and the results of this project, essentially all such concentrates
are and can be expected to be deficient in pyrite.   However, it is also
understood that all but a very few copper sulfide ore  bodies are charac-
terized by having an excess of pyrite present.  Suitable  copper sulfide-
iron sulfide concentrations for matte production  are almost  universally
obtained by depressing the flotation of the excess  pyrite and thus reject-
ing it to the tailings.  It is thus accepted that sufficient pyrite  for
the conduct of the proposed process already exists  in  almost all copper
sulfide ore bodies and that by moderate, tolerable  and probably desirable
changes in flotation process control, the required  amounts of pyrite could
be concurrently or separately recovered without difficulty or significant
economic penalty.  For chemical reasons as developed in this laboratory
study, separate recovery  and processing of the  required additional pyrite
appears preferable.
     The overall investment and operating costs as  determined by a prelimi-
nary economic study of the proposed copper smelting process  (through anode
production) with a capacity of 300 tons of copper per  day shows it to
involve an investment of  about $52,000,000 and  an overall smelting cost
for copper of about 12 cents per pound  without  credit  for recovered
elemental sulfur and iron oxide.  Both  of these resources are regarded as
marketable and at values  that are regarded as conservative under today's
conditions (at $20 and $10 per short ton for sulfur and  iron oxide,
respectively) the overall smelting cost for copper drops  to  about  9  cents.
Although a detailed estimate and comparison with  a  conventional smelter
was not made, it is understood that such smelting costs  (through anode

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                                    -5-
copper production) are 7 to 8 cents per pound,  but without equivalent
sulfur emissions control.   However, with such controls  another 5 to  6 cents
per pound would be added.   It is felt that a detailed study of the proposed
process could identify substantial  economies which would reduce the  overall
production cost significantly.
     The overall assessment of the  chemical, technical  and economic  feasi-
bility of the proposed process is definitely regarded as favorable at this
stage.
     Continued work at the bench-scale level to obtain data suitable for
pilot plant design, particularly work involving the continuous rather than
batch operation of the two principal processes of neutral roasting and acid
dissolution, is recommended.
     In view of the results of this study and the near and actual state-of-
the-art technology which may be employed for the principal processing
operations, it  is concluded that the overall proposed process has favorable
prospects for achieving an adequate and economically acceptable control
(97%) of sulfur dioxide emissions for conventional smelting processes
integrated with the proposed process.

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                                    -6-
                               INTRODUCTION

     Although in the U.S.  four copper producers* have  announced  plans  for
non-smelting type processes to augment normal  production  or  for  large  pilot
plant operations, a review of worldwide copper smelting practices  involving
about 100 major plants shows no trends to indicate a significant shift away
from conventional roaster-reverberatory-converter type overall operations.
Roasters, when required, are generally used to adjust  the sulfur content of
the charge to the reverberatory furnaces.  Reverberatory  furnaces  are  used
to adjust the iron content and to produce molten matte (a fluid  homogeneous
mixture of iron and copper sulfides).  Converters are  used to  produce
blister copper from matte.  Blister copper is  a relatively impure metal
containing some copper oxide, the precious metals and  other  impurities such
as nickel and selenium.  However, in order to  meet the increasingly more
stringent regulations on sulfur dioxide emissions, there  is  a  very definite
trend in the U.S. and elsewhere to initiate major changes in the smelting
practices, the objective being mainly to assure the production of higher
strength sulfur dioxide waste gases from which principally sulfuric acid
can be produced most efficiently and thus realize a major reduction in
sulfur dioxide emissions.   In many cases this  acid is  finding  use nearby
in expanded leaching operations for low-grade  oxidized copper  ores. The
changes in smelter practices do, however, involve major changes  in smelter
equipment such as for flash smelting and continuous converting.   One plant
in the U.S. (Tennessee Copper Co.) converts part of its waste  gas to liquid
sulfur dioxide.  The American Smelting and Refining Co. plant  at Tacoma,
Washington, is about to begin very large-scale production of liquid sulfur
dioxide from a portion of the concentrated converter gas  that  is not used
for sulfuric acid manufacture.  Although consideration is being  given  at
some plants to convert concentrated gas to elemental  sulfur or to dispose
* Cyprus Mines:  The Cymet Process
  Duval Corp.:  A Chloride Leach Process
  Anaconda:  A Low-Pressure Ammonia Leach Process
  Hecla:  A Roast-Leach-Electrowin Process

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                                    -7-
of excess sulfuric acid by conversion to waste gypsum,  no plants  are known
to be operating such processes on a large scale today in the U.S.   One new
copper producer (Hecla Mining Co.) is to operate roasters to produce oxide
copper and sulfuric acid for a leaching and electrowinning process.
     In general, the conversion of waste sulfur dioxide to elemental sulfur
has been regarded as an ideal solution to the problem.   Activity  in this
direction has generally been concerned with the chemical reduction of the
sulfur dioxide utilizing natural  gas or coke as the reductant.  At some
point in the chemistry of natural gas reduction, hydrogen sulfide is
regarded as being formed, then the usual reaction between hydrogen sulfide
and unconverted sulfur dioxide proceeds to form elemental sulfur.   Thus
hydrogen sulfide appears as a logical reagent for effective conversion of
sulfur dioxide to elemental sulfur.  Consideration of processes for hydro-
gen sulfide production without the use of a chemical  reductant  such as
natural gas have not received much attention.  However, this was  the objec-
tive of the earlier efforts on this present project where H2S production
was accomplished by reaction of high-temperature water vapor with the
residual iron sulfide obtained in an initial step of neutral roasting of
the pyritic copper concentrate.  This process was shown to be unfavorable
because of unsatisfactory equilibrium conditions from which only  very dilute
(less than 1%) hydrogen sulfide gas was obtained.  Accordingly, such pro-
duction and effective conversion  of the iron sulfide was achievable only by
treatment with very large volumes of high-temperature water vapor.  Pre-
liminary economic studies showed  this process to have unlikely  prospects
for economic viability.
     An alternative process was subsequently scoped.   In this process
(1) neutral roasting would still  be required (including in many cases pro-
vision of additional pyrite) followed by (2) acid treatment for hydrogen
sulfide production, (3) acid regeneration and recovery, and (4) enriched
copper sulfide recovery for delivery to conventional  converters.   The
sulfur dioxide from the converting step would be (5) reacted with an ade-
quate amount of hydrogen sulfide  from the acid treatment step to  produce

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                                    -8-
elemental sulfur.  This sulfur, together with that evolved from the neutral-
roasting step, would comprise essentially all of the sulfur present in the
original pyritic copper concentrates.   Thereby overall  production of ele-
mental sulfur would be achieved by thermal means without the need for a
specific chemical reductant.  This present report covers the details of
the laboratory efforts directed at assessing the chemical  feasibility of
such a process and its economic potential.  The principal  areas for study
include:  (1) the neutral-roasting step and (2) the acid-leaching or
hydrogen-sulfide production step.  Related matters concerned with feed,
residue and gas composition were studied concurrently.
     Most copper sulfide resources in the U.S. and throughout the world
are associated with pyrite.  For the recovery of concentrates from these
ores flotation treatment is generally used.  In this treatment excess
pyrite which is generally present in substantial amounts is discarded
along with the gangue.  This is accomplished by simple chemical adjustments
in the concentrator to preferentially depress the flotation of the iron
sulfide  (pyrite) and thus reject it from the circuit.  Since in most cases
additional iron sulfide is essential to the process to provide the required
amounts of hydrogen sulfide, it is a requirement of the suggested process
concept  that sufficient pyrite be retained with the copper sulfide concen-
trates  (or separately recovered) for subsequent use.  This appears to be
easy  to accommodate in existing concentrators.

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                                    -9-


                                BACKGROUND

THE OVERALL CONCEPT
     Although the individual  steps of the proposed process  concept comprise
well-known and fairly well-understood chemistry and technology,  the combi-
nation of these steps or processes to achieve the desired objective specif-
ically in copper smelting appears not to have been explored.   The patents
             /I ?\
by S. I. Levyv ' ' employ similar steps  specifically to treat pyrite for
recovery of iron, sulfur and other metals such as copper, lead and zinc.
These patents disclose the combination of steps of neutral  roasting, the
recovery of sulfur from neutral  roasting, hydrogen sulfide  conversion to
sulfur, hydrochloric acid treatment to preferentially dissolve iron and
liberate hydrogen sulfide, the separation of a copper-rich  residue, the
regeneration of hydrochloric acid with iron oxide recovery  and,  in addition,
the recovery of zinc and lead by separate electrolytic processes.  The
technology employed for iron and hydrochloric acid processing was different
from more recent processes by involving  crystallization of  the chlorides as
the primary means of separation from other metal  chlorides  which may be
present in substantial amounts.   The inventions do not include the further
processing of insoluble copper residue or the utilization of hydrogen sul-
fide to react with any sulfur dioxide derived therefrom.

NEUTRAL ROASTING
     The literature on the chemistry and technology of neutral roasting
wherein pyrite is thermally decomposed to yield elemental sulfur and
pyrrhotite is extensive and goes back many years; however,  much of it is
not of sufficient relevance to be useful in the present investigation.
The most recent and useful compilation of background information has been
reviewed in a report by Parsons and Ingraham.  '  The thermochemical
mineralogy of the copper-iron-sulfur system has been intensively studied
by many investigators.^    '  This work has included consideration of the
extremely complex and still imperfectly defined central portion of the

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                                    -10-
system.  Although simple chemical  equations  can  be  written  to  indicate  the
possible decomposition of the many copper-iron-sulfide  compounds   which
may be present in practical  pyritic copper concentrates to  yield  FeS,  Cu~S
and sulfur, it is clear that such  simplifications do  not actually occur
under conditions which are encountered in the present process  concept.
     As to the application of neutral  roasting on a practical  scale  in
recent years, perhaps the most noteworthy activity  has  been that  of  the
Outokumpu Company in Finland relative  to their flash  smelting  developments.
This process has been indicated to utilize a neutral-roasting  process  on
                                                                  (?6}
pyrite to produce elemental  sulfur and molten iron  sulfides (FeS).
This process has sufficient flexibility so that  either  neutral  or almost
any degree of oxidative roasting can be utilized.   It is uncertain at
present whether or not elemental sulfur production  via  neutral  roasting is
actually continuing today.
     Kunda and Mehta and coworkers^  '  ' have disclosed the details of a
proposed very large-scale process  for  pyrite neutral  roasting  in  Canada.
These references are most useful for their engineering  and  economic  details.
Watkinson and Germain^  ' of Noranda,  also in Canada, have  reported  useful
laboratory data on neutral roasting of pyrite intended  for  large-scale
applications.  No references to work specifically  related to practical
copper concentrate neutral roasting were uncovered  although Tkachenko,  et
al., and Isakova, et al.,  ~32' reported on studies  of the thermal  decom-
position of pyrite, chalcopyrite and bornite.
     Although specific references  to the neutral roasting of pyritic copper
concentrates were not uncovered, the related work  including actual large-
scale applications to pyrite alone, which is a significant  fraction  of the
neutral roasting processing as intended in the proposed process concept,
leads to the conclusion that neutral roasting as proposed here is largely
a present state-of-the-art technology.

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                                   -11-
LEACHING
     The leaching of processed pyritic minerals with acids  for preferential
dissolution of iron comprises a very old concept.   Parsons  and Ingrahanr
have also extensively reviewed this.  Their review indicates many concepts
which recognize the relative insolubility of copper in such treatment pro-
cesses.  Van Weert, Ingraham and Thornhill  and co-workers^   "  '  recently
reported on the leaching of nickeliferous pyrrhotites and matte with hydro-
chloric acid.  In their discussion  of this  work, regeneration of hydrochloric
acid from hydrated nickel chloride  crystals was indicated as a commercial
process.  It was also mentioned that such technology was being extended to
ferrous chloride.  However, references to operating processes which utilize
the chemistry of leaching were not  uncovered.  Leaching is, therefore,
perhaps the major part of the proposed overall process which is definitely
not present state-of-the-art technology.  However, it appears to be a
technology which is easy to develop to a practical scale and has been
worked on by others on a laboratory scale at least.

ACID REGENERATION
     Hydrochloric acid regeneration from corresponding acid-ferrous chloride
solutions is also an old concept and recently, within the last ten years,
has assumed considerable importance as the preferred and alternative means
of providing a regenerative process for use in the steel pickling industry
in the U.S.  This process regenerates the pickling acid (hydrochloric acid)
and produces concurrently iron oxide suitable for recycle in steel making.
Reeve and Ingrahanr ' have extensively reviewed this technology which
derives mainly from European developed practices.  Installations in the
U.S. include the Ruthner-Dravo process and more recently the Lurgi process.
The Turbulator Process, used in Europe and licensed in the  U.S. by Haveg
Industries, has apparently not yet been utilized in the U.S.  The latter
process is reported to be able to accommodate solutions as  low as 0.5% and
as high as 11% by weight in acid and up to 35% in dissolved ferrous chloride.
This is support for the expectation that such processes have very wide ranges

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                                    -12-
of flexibility for processing solutions of differing  composition with
essentially no practical  limits.   Acid regeneration and concurrent iron
oxide production are clearly regarded as present state-of-the-art processes.
The only problem here appears to  be one of scale-up.   Although pickle
liquor processing as practiced is of substantial size, mainstream ferrous
chloride decomposition and hydrochloric acid regeneration as  contemplated
in the proposed concept may be on a scale at least ten times  as large  as
in a single pickle liquor regeneration plant.   The selection  of which
existing process of those indicated above to use would be determined by
ease in scale-up and ultimate detailed construction and operating costs.

CONVERTING
     Converting of matte to blister copper is  a well-established state-of-
the-art process in essentially all sulfide copper smelters throughout  the
world.  In the proposed process concept, a feed richer in copper sulfide
would replace the usual matte composition.  It is understood  that such a
change can be accommodated by adjustments in fuel and air or  oxygen delivery
to the converter to compensate for the absence of the large amounts of iron
sulfide which are present in conventional mattes.  The lowered iron in the
feed will also result in substantial reductions in the amount of slag  which
will be produced by the converter.  However, special  (yet conventional)
means would probably be needed to prepare this relatively small amount of
slag for delivery back to the concentrator for flotation treatment for
residual copper recovery.

SULFUR PRODUCTION
     Sulfur production achieved by catalytically reacting sulfur dioxide
and hydrogen sulfide is an old state-of-the-art process.  Its principal
application today is the process  by which "sour gas"  (natural gas contain-
ing hydrogen sulfide, normally as a major contaminant) is purified for
widespread distribution and use.   The reaction is also employed commer-
cially in the petroleum refining  industry as the means of cleaning the

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                                   -13-
vent gases which may have intolerable hydrogen sulfide concentrations.   In
these processes (which are modifications of the so-called Claus  process),
the extracted hydrogen sulfide is partially burned to form the required
ratio of hydrogen sulfide to sulfur dioxide.   This mixture of gases is
then reacted on a bauxite catalyst to yield elemental sulfur as  either  a
gas, liquid or solid, depending on the temperature employed.
     The Claus process, though effective and economical as normally
operated, does not achieve complete reaction.  The vent gas may require
further treatment before release.  Alternative processes have therefore
been receiving increasing attention.
     Another process which is fairly advanced in development is the U.S.
                                 (}}
Bureau of Mines Citrate Process/   ;  It is primarily being considered  as
a means for removing sulfur dioxide from smelter flue gas  (and has also
been proposed for power plant flue gases) by absorption of the sulfur
dioxide in buffered aqueous sodium sulfite-sodium citrate solutions
followed by contacting with gaseous hydrogen sulfide  (separately prepared)
to yield elemental sulfur.  The  recovered sulfur is  intended to be par-
tially converted back  to hydrogen sulfide by chemical  reduction to provide
the  required amounts of hydrogen sulfide.  This process is being installed
by  USBM on a demonstration scale using a sulfur dioxide-bearing flue gas
stream from the  lead smelter of  the Bunker Hill Company at  Kellogg, Idaho.
This process is  of special interest because of its expected capability  for
effective  removal of sulfur dioxide from very dilute  gas  streams with
accompanying high yields and effective  conversion to  elemental sulfur.   In
these features  it may  have significant  advantages over the  conventional
Claus process  for application  in the  proposed process  concept.  In any  case,
the high  concentrations  of sulfur dioxide  expected from the converter  gas
stream and also the  very  high  concentrations  of hydrogen  sulfide produced
in  the leaching step  suggest  that either a  Claus  process  or a Citrate  pro-
cess may  be  utilized.   If  the  Citrate process  demonstrates  substantially
higher overall  recoveries  of  sulfur from dilute flue gases,  it may be  the
recommended  choice.   Its  full  demonstration,  which is expected  to  be
initiated this year,  is  awaited  with  interest.

-------
                                   -14-
                           PROCESS DESCRIPTION
     The objective of the proposed process is to produce hydrogen sulfide
by leaching neutral roasted copper concentrate containing iron sulfide
with hydrochloric acid.  The hydrogen sulfide is formed by the reaction of
the acid with iron sulfide.  Assuming the iron is present as FeS, the H,,S
is formed according to the reaction:
The hydrogen sulfide can then be used to convert the sulfur dioxide in the
smelter waste gas streams, particularly the converter gas, to elemental
sulfur via the reaction

          2H2S(g) + S% , - '  3S(s) + 2H2(g)
                                                                      (36 }
Either a modified Claus process or the Bureau of Mines Citrate process v
can be used to produce the elemental  sulfur from the H^S and SC^ .
     The schematic flow diagram for the proposed process is presented in
Figure 1.  The principal operations in the process are as follows:
 1.  Run-of-mill flotation concentrate is dried at 150-200C to remove
     water.
 2.  The dried concentrate is fired at 800-1 000C in a neutral or reducing
     atmosphere to remove the labile sulfur.  The sulfur is condensed and
     collected.  The roasting operation converts the bulk of the iron in
     the concentrate to a form that is soluble in hydrochloric acid.
 3.  The neutral roasted concentrate is leached in hydrochloric acid
     (4-5N) to dissolve the iron and convert an equivalent amount of the
     sulfur to hydrogen sulfide.  In this step the copper sul fides are
     relatively insoluble.
 4.  The slurry from the leach tank is filtered and the solid residue
     washed with water.
 5.  The solids, containing the bulk of the copper and a small amount of
     iron, are sent to  the converter for copper recovery.

-------
                                        SULFUR
                    t
s~
-^ i


FLOTATION
CONCENTRATE
HC1 '
HC1 RE
Fe203 
1
 j
EACH TAN< |
1
WASH 	
IATER j 1,
V m
-. LEACH SOLUTION rT| -,

f
WASH
FWATER
PER  A
COPPER-BEARING SOLIDS

ER
COPPER-BEARING
^ KhblUUL " ^
TO CONVERTER
FIGURE 1.   Schematic Flow Diagram for Acid Leaching Copper Concentrates

-------
                                   -16-
 6.   The cooled  leach  solution, plus wash, which contains the bulk of the
     iron  and  small amounts of copper, is treated with a fraction of the
     H?S.   This  precipitates  the copper and some trace impurities, such as
     silver, as  the sulfides.  The precipitate  is separated by filtration,
     washed, and sent  to  the  converter for copper recovery.
 7.   The leach solution,  which now consists of  hydrochloric acid and iron
     chlorides (principally as ferrous chloride), is  treated to convert
     the iron  chlorides to ferric oxide and recover   the hydrochloric acid
     for recycle.   Commercially available processes used for treating
     hydrochloric  acid pickle liquor can be used to process the leach
     solution.
 8.   Make-up  hydrochloric acid is added to the  process as necessary, to
     maintain  the acid inventory.
 9.   In the converter  the copper sulfide is reacted to form copper and SO,,.
10.   The  H?S  from the  leaching step  is collected and  used to reduce  the
     SCL  from the converter  to elemental sulfur.
11.   The  by-product sulfur  and ferric oxide can be  stockpiled  or  sold
     depending on market conditions.
     In order for the  process to achieve maximum utility  it  is  necessary  to
produce enough H2S in  the leaching  operation  to react with all  of  the  S02
produced in the copper smelter,  primarily  from the  converting  step.  This
means that about two-thirds  of the  sulfur  in  the neutral-roasted  concen-
trate must be converted to  H2S.   In actual  practice more  than  two-thirds
conversion must be obtained.   There are  two  reasons for  this:   (1)  some
H?S is needed to precipitate the dissolved copper  and trace  impurities in
the leach solution; and  (2)  some S02 is  formed in  the neutral-roasting
operation.  It may be required that this  S02  also  be  converted to elemental
sulfur.   In a typical  commercial  operation it would probably be necessary
to convert about 70% of  the sulfur in the neutral-roasted concentrate  to
H?S in order  to convert  all  of the S02 produced in the plant to elemental
sulfur.   For  concentrates which are deficient in iron sulfide, additional
amounts of iron sulfide  (pyrite) would have to be provided.   This is not
regarded  as a particularly critical problem because most sulfide copper

-------
                                    -17-
ore bodies contain excess pyrite which is rejected by conventional  control
in the concentrator's flotation operations.   Such excess  pyrite  could
therefore be separately recovered or allowed to accumulate with  the copper
sulfide concentrate.   It is probable that only a very few major  sulfide
ore bodies in the U.S. would be sufficiently deficient in pyrite to impair
the likelihood of realizing a practical  overall process.*
* The White Pine, Michigan, operation of the Copper Range Co. is the
  typical example.

-------
                                   -18-
                           EXPERIMENTAL PROGRAM

     The principal  operations in the proposed process  were demonstrated on
a laboratory scale using glass and Vycor reactors.   To simplify the experi-
mental  procedures the operations were carried in batch reactors rather
than continuously.

COPPER CONCENTRATES STUDIED
     In order for the proposed process to be applicable to a  smelter opera-
tion, enough H?S must be formed in the leaching reaction to combine with
all  of the S(L generated in other parts of the smelter (2 moles of HS
required per mole of S(L).   The composition of the  neutral-roasted concen-
trate can have a significant effect on the leaching reaction  and on the
HpS formed.   To evaluate the effect of composition, seven different copper
concentrates, as well as pyrite, were studied in this  program.   The con-
centrates were obtained from various domestic copper producers  as typical
run-of-mi11   flotation concentrates.  The chemical  and mineralogical com-
positions of the concentrates studied are given in  Table I.  (The mineral-
ogical  compositions are those reported by the suppliers of the  concentrates.)
Three of the concentrates (Morenci, Pima, and Anaconda) were  considered to
be typical of the general types of copper concentrates produced in the
United States and were studied in considerable detail.  The Morenci con-
centrate was obtained from the Phelps Dodge Corp. and  is principally a
mixture of chalcocite and pyrite.  The Pima concentrate was also obtained
from the Phelps Dodge Corp. and is a mixture of chalcopyrite  and pyrite.
The Anaconda concentrate was obtained from the Anaconda Company and is a
complex mixture of iron and copper sulfides as well as zinc,  arsenic and
antimony sulfides.
     All of the concentrates contain varying levels of minor  constituents
which can affect their use in the proposed process.  These include those
materials which can affect the safety aspects of the process  such as
arsenic, antimony, mercury, etc., and the precious  metals, gold and silver,

-------
                                  -19-
           TABLE I.  Composition of  Concentrates Evaluated
                     (Chemical  Composition  is  in Weight
                     Percent)
1




Element





Cu
Fe
S
Insol.
As

Sb
Zn

Pb
Mo
Bi
Cd

Se
Te
Hg
Ag*
Au*

















>P_
y
C

O 3 T-
0 r- S-
i 0. >,
i 3 d.
i .C
: O
	




(0
E

O-


27.14
27.3
29.6
10.16
0.06

0.025
0.30

0.04
0.02
0.016
0.002

0.038
0.0042
0.028
2.8
0.01


Ol

 r-
t- Ol
>-> 1/1 +J
D. 3 T-
Or- 1-
U 0- >i
i Q.
fO
JC
o
" -*~  

IO
x>
C
o
u

C



26.93
19.4
34.7
6.64
1.86

0.037
6.2

0.05
0.05
0.03
0.12

0.029
0.025
0.024
10.42
0.07

3"
_Q
"I/I I/)
01 01
Li_ ""O
C -r-
X M 4-
QJ r
r -3
o. tn in
E <
o
o




01
C
o

>-,
1


20.31
30.1
40.0
6.48
0.39

0.05
0.3

0.04
0.085
0.012
0.001

0.028
0.0058
0.031
1.54
0.01




Ol
^J
r- Q)
O 1/1 4J
O 3 -r-
U f l-
1 0. >,
1X3 Q.
JC
O


i.
0>
-o
c +->
Ol -r-
> CI-
IO
_ J


12.65
32.0
39.7
8.88
0.36

0.03
1.8

0.20
0.025
0.024
0.004

0.036
0.011
0.024
1.94
0.07

CJ

* t/1
O> Ol
u. -o

X 1-
01 

CL tn
E
o



C
 f
CU fO
4J C
-(-> 3
to O
CQ s:


	
25.0
31.3
13.12
0.06

0.063
0.70

0.03
0.018
0.020
0.003

0.051
0.0064
0.038
10.79
0.31


. "

O) *r- OJ
4J &~ 1 *
r- >,-r-
0 0. i-
0 0 >,
U U O-
r i 
IT5 ID T3
-C .C c:
O U IO
) 	
Ol
3
C
IO

C
<0
CO
"

29.58
27.28
29.49 '


















01

'r-
i.
>^
D.
O
u
i 
0
.C






IO
u
01
1C.



36.96
44.23
























Ol
+J
r"
L.

O.
In  troy ounces per ton

-------
                                   -20-
which must be recovered for their economic value.  Therefore, it is neces-
sary to know the fate of the minor constituents as the concentrates are
processed.
     Pima and San Manuel materials are typical of chalcopyrite concentrates.
Morenci and Tyrone materials are typical of chalcocite concentrates.
Anaconda and Lavender Pit materials are typical of complex sulfides charac-
teristically high in zinc and arsenic.  Battle Mt. and Anaconda concentrates
are characteristically high in precious metals, gold and silver.
     If one ignores the minor constituents, the concentrates have the
following compositions:
          Morenci           Fe0.87Cu0.58S2     (M0.73S)
          Pima              Fei.o6Cu0.92Ss     (M0.99S)
          Anaconda          Feo.64Cu0.78S2     (M0.72S)
          Tyrone*           Fe0.86Cu0.51S2     (M0.69S)
          Lavender  Pit*     Fe0.93Cu0.32S2     (M0.63S)
          Battle Mt.**      Fe0.99CV85S2     (M0.92S)
          San Manuel***     Fei.Q6Cul,01S2     ^M1.04S)
          Hecla****        FeS2.09
    * Also obtained from Phelps Dodge Corporation representative of their
      Tyrone, New Mexico, and Bisbee, Arizona,  operations.
   ** Obtained from the Duval Corporation's Nevada operation.
  *** Obtained from Newmont Mining Company's Magma Copper,  Arizona,
      operations.
 **** pyrite concentrate representative of material from Utah, obtained
      from Hecla Mining Co.

-------
                                   -21-
NEUTRAL ROASTING OF CONCENTRATES
     The neutral roasting of the various concentrates was carried out in an
atmosphere of purified argon or helium, or under a vacuum (<100 microns
absolute pressure).  The concentrates were roasted at temperatures from
800 to 1020C.  A Vycor reactor was used to contain the concentrates at
roasting temperatures up to 900C; alumina was used for temperatures
above 900C.  The general procedure used for neutral  roasting the concen-
trates was as follows.
          A weighed amount of previously dried concentrate was placed
     in a horizontal Vycor reactor tube; the tube was purged with
     argon or helium.  The reactor was heated in a tube furnace to the
     desired temperature and maintained at temperature for 6-24 hours.
     The reactor was then cooled to room temperature.  The inert gas
     flow was maintained throughout the cycle.  After cooling, the
     neutral-roasted concentrate was removed from the tube and ball-
     milled for one hour in a porcelain mill using alundum balls.  The
     milled concentrate was screened through a 70-mesh screen and
     stored in a sealed glass jar.  Oversize material was crushed
     and milled again.
     When the neutral roasting was carried out in an alumina crucible, the
procedure was modified as follows.
          A weighed amount of concentrate was placed in an impervious
     alumina crucible and the crucible was placed in a closed-end
     Vycor tube.  The tube was placed in a pot furnace and purged
     with argon.  The concentrate was then roasted in the manner
     described above.
     When the roasting was carried out in a vacuum, a two-stage vacuum pump
was used to maintain the vacuum.  Two traps were placed between the reactor
and the pump to prevent SO,, and sulfur from entering the pump.
     The bulk of the concentrates were roasted at 800C under argon in
one-kilogram batches.  In the case of the Morenci, Pima, and Anaconda

-------
                                   -22-
concentrates, several  one-kilogram batches of each were neutral-roasted at
800C.  The various one-kilogram batches of each neutral-roasted product
were combined and blended by ball  milling for one hour.  The composite
blend of each concentrate was used for the bulk of the leaching studies.
     Smaller batches of each of the three concentrates were neutral-roasted
at temperatures above 800C to determine the effect of temperature on the
product and on the leaching reaction.  At the maximum roasting temperature
HOOOC) the three concentrates were partially melted.

ACID LEACHING OF NEUTRAL-ROASTED CONCENTRATES
     Leaching of the neutral-roasted concentrates in hydrochloric acid was
carried out in the laboratory employing equipment as illustrated in Fig-
ure 2.  The reaction vessel is a one-liter three-necked flask.  The pro-
cedure used in the leach tests was as follows.
 1.  The reaction flask was placed in a thermostatically controlled water
     bath.
 2.  The required volume of hydrochloric acid was added to the reaction
     flask.
 3.  The water in the bath was heated to the required  temperature using a
     cartridge heater and  thermoregulator.  The temperature of the hydro-
     chloric acid was monitored with a  thermometer.  The bath temperature
     and acid temperature, at equilibrium, never  varied by more than  1C.
 4.  Water  in the  bath and acid in the  reaction flask  were stirred using
     magnetic stirring bars and a  stirring hot  plate.
 5.  While  the acid was  being heated the  reactor  was purged with argon.
      Off-gas from  the reactor passed through a  water-cooled condenser and
      two  traps.  The  bulk  of  the water  in  the off-gas  was condensed  in the
      condenser and  returned to  the reaction  vessel.  The first  trap  con-
      tained a dilute  (^.BH)  hydrochloric  acid  solution and served to
      collect any hydrogen  chloride in  the  off-gas.   The second  trap  con-
      tained a  known  volume (usually  1  liter)  of 1.5N. sodium hydroxide and
      was  used  to absorb  the  H^S in the  off-gas.

-------
       CONDENSER
COOLING WATER

        ARGON

      TO POWER
       SUPPLY
THERMOREGULATOR 
    HEATER
GLASS TANK


  WATER
                       I
                                WATER
                                                ACID  *.
                                              SCRUBBER
                                              0.5N  HC1-
                               THERMOMETER
                                 1-LITER 3-NECK
                                     FLASK
                                     HYDROCHLORIC ACID
                                       + CONCENTRATE


                                      MAGNETIC
                                      STIRRERS
                   STIRRING
                   HOT  PLATE
                                                                           H?S  ABSORPTION
                                                                                   .5M  NaOH


                                                                                   MAGNETIC
                                                                                   STIRRER
                                                                  STIRRING
                                                                  HOT PLATE
       FIGURE 2.  Equipment for Leaching Neutral-Roasted Concentrates in Hydrochloric Acid

-------
                                   -24-
 6.   When the acid  reached  the  desired  temperature  the concentrate was
     added.
 7.   The course  of  the  reaction was  followed  by  absorbing the H^S in the
     caustic solution  (designated  as scrubber solution or off-gas absorber
     solution) and  analyzing  the solution  periodically for  its  sulfide
     content.
 8.   In some runs the  solution  in  the reaction vessel was sampled periodi-
     cally,  filtered,  and  analyzed for iron and  copper.
 9.   The run was normally  continued  until  the sulfide concentration  in  the
     scrubber solution  (step  7  above) reached a  constant  level  indicating
     dissolution of the iron  in the  acid was  complete.
10.   When the run was  complete, the  reaction  slurry was filtered and the
     solid residue  washed  with  water.
11.   The solid residue was  dried at  110C, weighed  and analyzed for  iron,
     copper  and  sulfur.  (In  certain runs  the residue was also  analyzed
     for minor constituents.)
12.   The filtrate and  wash solutions were  combined, diluted to  a known
     volume  and  analyzed for  iron  and copper.  (As  in 11  above  in certain
     runs the solution was analyzed  for minor constituents.)
13.   Material balances for the  iron, copper and  sulfur were then calculated
     based on the feed, solid residue, leach  solution, and  scrubber  solu-
     tion analyses.
     After several  runs it became  apparent that  the amount  of  H^S liberated
was a good measure of  the  iron  dissolved (normally  very  little  copper  dis-
solved) and  was  the most reliable  method of following the leaching  reaction.
For this reason in most of the  runs  no attempt was  made  to  sample and
analyze the  leach solution during  the course  of  the run.   Only the  final
solution was analyzed  for iron  and copper.
     Some runs were made at the boiling point of the hydrochloric acid. For
4N HC1 this  was about 106C.    Since temperature control  was not  critical,
as long as the solution was boiling, the water bath was  not needed.   Instead
the reaction vessel was heated directly using the stirring  hot plate.

-------
                                   -25-
ANALYTICAL PROCEDURES
     A variety of procedures was used to analyze the concentrates,  neutral -
roasted concentrates, leach solutions, leach residues and scrubber  solu-
tions.  In addition, considerable use was made of a commercial  assayer to
analyze the concentrates, neutral-roasted concentrates and leach residues.
Results obtained from the assayer were checked periodically by  in-house
analysis of duplicate samples.  The assayer was employed for economic
reasons.  In-house analytical facilities could not compete on a cost basis
with the assayer on routine analyses such as iron, copper and sulfur in
concentrates.
     The H~S evolved during the leaching was determined by absorbing the
H^S in sodium hydroxide solution and analyzing the solution for sulfide
ion.
          H2S + 2NaOH  	> 2Na+ + S= + 2H20

The sulfide content of the caustic solution was determined using a  sulfide
specific ion electrode (Orion Model 94-16A) and calomel reference electrode
(Beckman Perma-Probe solid state reference electrode Model 39407).   The
following procedure was used.
          A known volume of the sulfide-caustic solution (usually
     2-5 mis) was added to about 150 ml of 1M_ sodium hydroxide  solu-
     tion.  The solution was thoroughly mixed using a magnetic
     stirrer.  The sulfide sample was then titrated with 0.1^ silver
     nitrate solution.  The titration end point was determined  using
     the sulfide electrode, reference electrode and a digital milli-
     volt meter.
A typical titration curve is shown in Figure 3.  The point of greatest
inflection is taken as the end point.  The procedure precision  was  found
to be 2.0%.
     Iron in the leach solutions was determined using the procedure
developed by Harvey, et al/37' and a Gary Model 14H recording spectro-
photometer.  The procedure allows the simultaneous determination of both

-------
                               -26-
  -1000
   -800
   -600
   -400
^  -200
    200
    400
    600
                                    END  POINT
5ml SAMPLE-NaOH SOLUTION
  I        I	I	I
                         8       12      16
                         0.1M  AqN03 ADDED,
                                  20
24
   FIGURE 3.   Typical  Sulfide Titration Curve Using  Sulfide Electrode

-------
                                    -27-
iron (II) and total iron in the solution.  Iron (III) is obtained by
difference.  Impurities present in the leach solution, such as copper and
chloride, do not interfere at the levels present.   The procedure used is
as follows:
          A known volume of leach solution (50-100 lambda) is placed in a
     50 ml volumetric flask.  Ten ml  of 0.3% 1,10 phenanthroline and 5 ml
     of 0.2M potassium acid phthalate are added.   Distilled water is added
     to the mark and the solution stirred.  The absorbance of the solution
     is read (within 30 minutes) on the Gary at 396 my and 512 my.   The
     precision of the procedure is about 1.5%.
Preparation of the standard concentration curves  is rather complex as is
calculating the iron (II) and iron (III) from the curves.  The reader is
referred to the original reference^  ' for complete details on the
procedure.
     Copper in the leach solutions was determined by atomic adsorption
using standard procedures.
     The procedures described above were also used to determine iron and
copper in the concentrates, neutral-roasted concentrates and leach residues
after they had been solubilized by fusion techniques.  Sulfur in the solids
was determined using the Leco S0~ analyzer.
     The various concentrates and neutral-roasted concentrates were analyzed
by thermal techniques using the DuPont Model 990 Thermal Analyzer and
Model 951 Thermogravimetric Analyzer.

-------
                                   -28-
                          RESULTS AND DISCUSSION

     The basic steps in the proposed process were demonstrated on a labora-
tory scale.  These included the neutral  roasting, acid leaching,  and copper
recovery from the leach solution.  Conversion of the ferrous chloride to
ferric oxide and recovery of the hydrochloric acid was not studied in detail
because commercial processes are available for this operation.  To simplify
the equipment required batch reactors were used for the neutral roasting
and leaching operations.  On a commercial  scale, however, both operations
would expect to be carried out in continuous reaction systems.

NEUTRAL ROASTING
     Neutral roasting of the copper concentrates is necessary in order to
carry out the acid leaching operation.  When run-of-mill flotation concen-
trate is leached with hydrochloric acid there is very little reaction and
no production of H2S.  When the concentrate (or pyrite) is heated to an
elevated temperature (>500C) in an inert atmosphere part of the sulfur is
evolved and the residue obtained is partially soluble in hydrochloric acid.
Almost all of the iron in the neutral-roasted concentrate dissolves forming
ferrous chloride with the evolution of H2S.  The neutral-roasting operation
is, therefore, a vital part of the overall process.
     Substantial volumes of each concentrate were roasted under various
conditions for use in the leaching studies.  In addition, thermal analysis
(DTA and TGA) was used to determine the optimum conditions for the neutral-
roasting operation.  Chemical and X-ray analysis were used to assess the
composition of the neutral-roasted concentrates.
Composition of Neutral-Roasted Concentrates
     When  the concentrates are heated in an inert atmosphere  (or vacuum)
water and  elemental  sulfur  (plus small amounts of H2S and S02) are evolved.
In a static system the amount of sulfur released will depend  on the min-
eralogical  composition,  system temperature and the  sulfur partial  pressure

-------
                                    -29-
(assuming equilibrium is attained).  In a flowing system, where the sulfur
is removed as it is formed, the amount of sulfur evolved will depend on the
mineralogical composition, the reaction temperature and the time at tempera-
ture.  The geometry of the concentrate bed can affect the efficiency of
sulfur removal and thus affect the time required to achieve a given sulfur
loss.  This will be discussed in a later section.
     For the leaching studies, the bulk of the concentrates were neutral-
roasted at 800C in flowing argon or helium (no differences in product
composition were detected as a result of interchanging the inert gases).
They were processed in one-kilogram batches.   Where more than one kilogram
of a given concentrate had to be treated, several batches were processed
and the products combined and blended by ball  milling.  The compositions
of the neutral-roasted concentrates, prepared  at 800C, are presented in
Table II.  It is impossible to assign specific formulas to the neutral-
roasted concentrates.  However, ignoring the minor constituents, the
neutral-roasted products have composition ratios as listed following Table II

          TABLE II.  Composition of Neutral Roasted Concentrates
Concentrate
i
Morenci
Pi ma
Anaconda
Tyrone
Battle Mtn
1 Lavender Pit
I
San Manuel
Hecla
I (pyrite)
Temp.
TO
800
800
800
800
800
800
800
800
Atmosphere
Argon
Argon
Argon
Helium
He 1 i urn
Helium
He 1 i urn
Helium
Time at
Temp.
(hrs)
24
24
24
16
16
16
24
24
Product Composition, wt%
Fe
33.4
29.4
23.2
36.4
28.1
39.1
28.5
52.9
Cu.
24.7
29.4
34.9
23.9
28.1
15.4
32.5

S
28.8
27.8
26.5
28.6
24.9
29.0
26.7
33.9

-------
                                    -30-
          Concentrate         Composition          Fe+Cu/S

          Morenci             Fel.33Cu0.86S         (M1.10S)
          Pima               Fe-j  21Cu1.07S2        ^M1.14S^
          Anaconda           Fe-,  n-|Cu-|  ^2        ^1 17^

          Tyrone             Fel.46Cu0.84S2        (M1.15S)
          Battle Mtn         Fe1  2gCu1  ]4$2        (M1.22S^

          Lavender Pit       Fel.55Cu0.54S2        (M1.05S)
          San Manuel         Fe-,  09^ui  o-jS?        (^i ?^)
                               \  C.L.  I . .3            I  ^--J
          Hecla              FeS1  12

Comparing these compositions with those for the concentrates reported
earlier one can see how the metal-sulfur ratio changes due to sulfur loss
during neutral roasting.
                                                   Neutral Roasted
                              Concentrate             Concentrate

          Morenci               MQ 73S                 Ml.10S

          Pima                  M0.99S                 M1.14S
          Anaconda              Mg y2$                 ^1.17^
          Tyrone                MQ>6gS                 M]J5S
          Battle  Mtn            MQ g2S                 M1.22S
          Lavender  Pit          V\Q^S                 M^^S
          San  Manuel            M^ ^^S                 ^1.23^

          Hecla                 FeS2.09                FeSl.l2

      By  neutral  roasting at  temperatures  above 800C sulfur  removal  can be
 increased.   Table III  shows  the composition of Morenci concentrate which
 was  neutral  roasted at various temperatures.   While  there are  some varia-
 tions it is  apparent that  high temperatures and/or long  reaction  times

-------
                                 -31-
        TABLE III.   Effect of Temperature on Composition
                    of Neutral  Roasted Morenci  Concentrate
Sample
No.

1
2

3
4*
5**
Temp.
C

800
800

860
970
1005
Time
(hrs)

24
16

20
3
20
Atmosphere
Product Composition, wt% 
Fe

Argon
Argon

Helium
Vacuum
Vacuum
33.4
32.3

32.1
37.3
33.5
Cu
s i
!
24.7 > 28.8 !
!
23.7

24.6
28.5
29.2 i

27.0 1
30.5 j
25.9 j . 25.2 i
I 1
     Ignoring the minor constituents the samples have the
     following compositions:
     Sample No.
         1
         2
         3
         4
         5
 Composition
Fe1.33Cu0.86S2
Fe1.27Cu0.82S2
Fe1.37CY92S2
Fe1.40Cu0.94S2
Fe1.42Cu1.04S2
Fe+Cu/S
M1.10S
M1.05S
M1.15S
M1.17S
M1.23S
 * Product consisted of two phases:   a solid mass (^ 40 wt%) which
   had fused and a sintered mass.  The analysis given is for the
   fused portion.
** Product had fused.

-------
                                     -32-
increase the metal-to-sulfur ratio  of  the  neutral-roasted concentrates.
Similar data for neutral-roasted  Pima  and  Anaconda concentrates are given
in Table IV.

        TABLE IV.  Composition of Neutral  Roasted  Pima and Anaconda
                   Concentrates  Prepared at Various Temperatures
Concentrate
Pima
i
Pima* A
B
i
Pima
Pima
Anaconda
Anaconda
Temp.
(0
800
860
860
970
1005
800
1000
Time
(hrs)
24
72
' 72
18
6
24
4
Atmosphere
Argon
Vacuum
Vacuum
Vacuum
Vacuum
Argon
Vacuum
Product Composition, wt%
Fe
29.4
28.9
29.1
28.9
29.1
23.2
28.1
Cu
29.4
28.6
30.9
29.7
29.3
34.9
42.3
S
27.8
24,6
25.5
24.7
24.8
26.5
26.9
M/S
1.14S
M1.25S
M1.27S
M1.28S
M1.27S
M1.17S
M1.39S
 * Product consisted of two phases:   (A) a partially sintered easily crushed powder,
   and (B) a highly sintered hard mass.  The phases were separated and analyzed
   individually.
     The physical  characteristics of the neutral roasted concentrates  varied
with the reaction  temperature and time at temperature.  Concentrate  processed
at 800C consisted of partially sintered easily crushed lumps  and  powder.
Those  heated  to  1000C or higher had partially fused and were  more difficult
to crush and  grind.   Concentrates processed at intermediate  temperatures
varied from  loosely sintered powders to fused lumps depending  on  the time
and temperature.   The Pima concentrate showed a somewhat greater  tendency
to sinter  than  the others.

-------
                                   -33-
     The particle size distribution of the  Morenci  concentrate,  which  had
been roasted at 800C for 24 hours in argon and  then ball  milled,  was
determined using U.S. Standard Sieve Series Screens.  The  distribution was
as shown in the following:
                Screen Size               Cumulative % Retained
                   +70                            0.08
                   +100                            6.04
                   +140                           21.78
                   +200                           47.84
                   +230                           72.35
                   +325                           89.96
                   Total                         100.00
The material had a tap density of 2.34 g/cm3, and the surface area was
<0.5 sq m/gram.
X-Ray Analysis of Neutral-Roasted Concentrates
     The various neutral-roasted concentrates were analyzed by X-ray dif-
fraction in an attempt to identify the minerals present.  These attempts
met with only marginal success and it was impossible to identify, with
certainty,  the minerals present in the neutral roasted concentrates.  A
detailed study of the mineralogical composition of  the neutral-roasted
concentrates was beyond the scope of this program.
     The iron-copper-sulfur system is extremely complex.  Despite decades
of  investigation by many different workers the phase diagram for the  system
has not yet been completely defined.(5~25'  Table V is a list of most of
the minerals  (stable  at low temperatures) reported  in  the literature  for
the Fe-Cu-S system.^5'   It does not  include questionable or incompletely
described  minerals which  have been reported.  Figure 4 shows the Fe-Cu-S
system  and location  in  the diagram of the various minerals listed in
Table V.   The  locations of  the neutral-roasted concentrates  (ignoring  the
minor constituents)  are also  indicated.

-------
                                -34-
           TABLE V.  Compounds in the Fe-Cu-S System that
                     are Stable at Low Temperatures(5)
Symbol
 a-bn
 an
 bbcv
 bn
 cc
 cf
 cv
 cb
 di
 dj
 fk
 gr
 he
 h-po
 Id
 mk
 ma
 mh
 m-po
 py
 sm
 tal
 tr
          Name
anomalous bornite
an i1i te
blaubleibender covellite
bornite
chalcocite
chalcopyrite
covellite
cubanite
digenite
djurleite
fukuchilite
greigite
haycockite
hexagonal pyrrhotite
Idaite
mackinawite
marcasite
mooihoekite
monoclinic pyrrhotite
pyrite
smythite
talnakhite
troilite
probable new mineral
new mineral
synthetic mineral
    Formula
 Cu5FeS4
 C1.75S
 CU1.1S
 Cu5FeS4
 Cu2S
 CuFeS2
 CuS
CugS5
CU1.96S
Cu3FeS8
Fe3S4
Fe9S10
Fe1.06S
Fe7S8
FeS2
Fe3S4' Fe3.34S4
Cu9Fe8S16
FeS
CucFeSc
  D   D
Cu0.12Fe0.94S
Cu3Fe4S6

-------
                                         -35-
Cu
                                                                              Fe
               FIGURE  4.   Mineral  Compositions  in  the  Fe-Cu-S  System
                          (See  Table  V for meaning of  symbols.)

-------
                                   -36-
     The compositions of the neutral  roasted  concentrates  (except  for  the
                                                              (5)
Hecla) fall  in the central  area of the Fe-Cu-S system.   Cabri,v    Yund and
Kullerud^24' and others have shown that at elevated  temperatures  the cen-
tral  area of the Fe-Cu-S system is characterized by  a  large solid  solution
field.  The composition of the neutral roasted copper  concentrates place
them within the solid solution field at temperatures of 700C and  above.
According to Cabri^ "The central area of the Cu-Fe-S system is  charac-
terized by a large solid solution field at elevated  temperatures  which
breaks up into five distinct phases at low temperatures.  The picture  is
complicated by numerous phase transformations, unquenchable phases, and
phases with closely related crystal chemistry, resulting in their having
similar physical appearance and X-ray diffraction powder patterns.  The
slow rate of most of the reactions at low temperatures also makes it diffi-
cult  (in some cases impossible) to achieve equilibrium in the laboratory."
This statement helps explain the difficulties encountered in trying to
identify phases present in the neutral roasted concentrates.  When the con-
centrates were neutral-roasted they were held at temperature for limited
periods of  time (4-24 hours).  It  is  not known if this time was sufficient
for solid solution to occur.   In addition, no attempts were made to control
the quench  rate when the concentrates were cooled.  Therefore samples of a
given concentrate which were neutral-roasted under approximately similar
conditions  could  have different mineralogical compositions depending on
such  factors as variations  in  cooling rate.  For example, four samples  of
air dried Morenci concentrate  were heated to 800C in flowing helium on the
thermobalance and then  cooled  to  room temperature.  The cooling rate used
was the  normal cooling  rate of the thermobalance furnace with the power off
(approximately  90 minutes  from 800 to 50C).  The samples were then analyzed
by X-ray diffraction.   The  diffraction patterns for the four  samples are
presented in  Table VI.   It  is  readily apparent  that there are significant
differences between  the patterns  as well  as  great similarity.  The lines of
highest  intensity are  very  similar for all four samples.   Similar  results
were  observed with the  Pima and  Anaconda  concentrates.

-------
                              -37-
TABLE VI.  The X-Ray Diffraction Patterns for Samples of
Morenci
at 800C
Sample
No. SM-1
dA
6.46
3.37
3.28
3.16
3.09
3.01
2.75
2.67
2.51
2.37
2.09
1.97
1.94
1.93
1.89
1.73
1.64
1.61
1.33
1.22
1.11
1.09
1.03
I/Io
15
20
25 .
25
100
20
25
30
15
20
35
20
35
40
70
20
20
35
20
15
15
20
20
Concentrates which were Neutral Roasted
in Flowing Argon on the Thermobalance
Sample
No. SM-2
dA
6.50
4.56
3.36
3.15
3.09
3.01
2.90
2.78
2.66
2.08
1.93
1.89
1.81
1.73
1.61
1.57
1.33






I/Io
15
15
10
100
55
10
10
10
20
25
20
30
10
15
15
10
5






Sample
No. SM-3
dA
6.50
3.37
3.30
3.15
3.11
3.00
2.75
2.67
2.54
2.10
1.93
1.89
1.75
1.61
1.33
1.15
1.11
1.09





I/Io
30
20
25
45
100
25
20
30
15
45
50
55
20
30
20
15
15
15





Sample
No. SM-4
dA
6.42
3.36
3.11
2.68
2.10
1.89
1.61
1.45
1.33
1.23
1.09
1.03











I/Io
10
15
100
10
10
55
20
5
5
10
10
5












-------
                                   -38-
     Samples of neutral-roasted Pima  which  were  prepared  at  different  tem-
peratures exhibited completely different X-ray diffraction patterns  (see
Table VII).   Samples of neutral-roasted  Anaconda prepared at different
temperatures also showed a lack of similarity  in their  diffraction patterns,
In the case of neutral-roasted Morenci  the  situation  was  different.  The
samples prepared at different temperatures  had similar  but not  identical
diffraction patterns.
   TABLE VII.  X-Ray Diffraction Patterns for Neutral  Roasted Pima
               Concentrate Prepared Under Various Conditions
            Sample P.-1*                     Sample P-8**
          dA         I/Io
          3.40         15
          3.10        100
          2.67         15
          1.89         80
          1.61         30
          1.55          5
          1.34         10
          1.33         10
          1.31         10
          1.22          5
          1.09         10
dA
6.31
3.30
3.15
3.10
3.00
2.72
2.67
2.53
2.10
1.93
1.75
1.73
1.64
1.33
1.09
I/Io
30
30
60
30
40
30
40
10
85
100
20
35
20
5
5
  * Concentrate neutral roasted for 24 hours at 800C
    in inert atmosphere.
 ** Concentrate neutral roasted for 6 hours at 1005C
    under vacuum.

-------
                                    -39-
      It was not possible to identify, with complete confidence, specific
mineral phases in any of the neutral roasted concentrates.  None of the
X-ray diffraction patterns obtained could be attributed specifically to
known patterns published in the ASTM Powder Diffraction File.
      In the case of the neutral-roasted Morenci and Pima concentrates
prepared at 800C their diffraction patterns were similar, but not identi-
cal, to those reported for mooihoekite (CuQFeQS,c) and talnakhite
            (5,7,8,10)
                  9 C9- 16'
The two materials probably contain either or both
minerals.  It is also possible they may contain some cubanite (CuFe2$3).
     In the case of neutral roasted Anaconda concentrate prepared at 800C,
the X-ray diffraction pattern could not be identified with any published
patterns for the Fe-Cu-S system.  The material prepared at 1000C had a
distinctly different pattern.  The pattern had a slight similarity to the
patterns for cubanite (CuPe^S.,) and bornite (Cu5FeS.), and the material
could be a mixture of the two minerals.
     The bulk of the leaching tests were carried out with concentrates
which were neutral-roasted at 800C.  In the case of the Morenci, Pima and
Anaconda, several batches of each concentrate were roasted at 800C and
blended.  The blended lot of each concentrate was used for the leaching
tests.   Therefore, differences in mineral  composition from batch to batch
due to slight uncontrolled variations in the neutral roasting operation
would be masked by the blending operation.
Thermal Analysis of Concentrates
     The neutral roasting of concentrates  was studied using differential
thermal analysis and thermogravimetric analysis.  Thermograms were obtained
for the various concentrates in purified helium.  Figure 5 shows weight
loss as a function of sample temperature for Morenci, Pima, and Anaconda
concentrates when heated in helium.  The thermograms for the other concen-
trates  are shown in Figure 6.  The initial weight loss (up to a temperature
of about 350C) is due to the  vaporization of water.  At higher temperatures

-------
  100






   96


i

o

UJ
s  92

i
<
z

CD


  88

U_
O




K--  84
Z
O

UJ
3

uj  80
_i
(X
s:




   76
72
         ATMOSPHERE  - PURIFIED HELIUM

         HEATING  RATE - 20C/MINUTE

         SAMPLE PREDRIED AT 110C
                                                                                 I
         100      200     300     400     500     600     700     800      900     1000


                                  SAMPLE TEMPERATURE, C
                                                                                                       o
                                                                                                       i
   FIGURE 5.  Thermogravimetric Analysis of Copper  Concentrates in  Helium

-------
   100
    96
    92
s  
o
    84
    80
    76
    72
                                                                 SAN MANUEL
                                                                           BATTLE  MTN.
                                                                            -_
          ATMOSPHERE-PURIFIED HELIUM
          HEATING RATE-20C/MINUTE

              I	I	I
I
J	I
I
             100     200     300     400      500      600     700     800     900    1000

                                      SAMPLE  TEMPER/JURE, C
                                                                                                         -p.
                                                                                                          i
                 FIGURE 6.  Thermograms for Concentrates in Helium

-------
                                   -42-
the weight loss is due to the evolution of sulfur (and  small  amounts  of
H?S and SCL).  In the case of the Anaconda concentrate  a  fraction  of  the
weight loss is due to vaporization of arsenic  sulfide and possibly zinc
sulfide.
     The data presented in Figures 5 and 6 were obtained  at a heating rate
of 20C/minute.  To show the effect of heating rate on  the decomposition
reaction(s) samples of Morenci  concentrate were heated  at various  rates.
The results obtained are presented in Figure 7.
     To obtain a better measure of the kinetics of the  decomposition
reaction samples of Morenci concentrate were inserted into a preheated
furnace and the weight loss followed as a function of time.  The thermo-
grams obtained are shown in Figure 8.  It normally required from 1 to
2 minutes for the sample to reach the reaction temperature.  During this
heat-up time, considerable sulfur evolution occurred, especially when the
control temperature was above 500C.  It was impossible,  therefore, to
measure precisely the evolution of sulfur as a function of time at tempera-
ture.  In obtaining the data presented in Figure 8 the  sample temperature
was monitored as a function of time as well as sample weight.  The dashed
portions of the thermograms represent the periods in which sample tempera-
tures were increasing to the control temperatures.  It  can be seen from
the data that a substantial fraction of the weight loss occurred before
the control temperature was reached.  At temperatures of 600C and above
most of the weight loss occurs in the first 25-50 minutes.  Thereafter the
loss of weight continues at a slow, almost constant, rate for long periods
of time.  One run at 800C, using Morenci concentrate,  was continued for
24 hours; a continual weight loss was observed throughout the run.  The
overall weight loss at the end of the 24 hour period was 26.06%.  The weight
change observed during the last few hours of the run was so slight, however,
it may have been due to instrument instability.
     The thermobalance data were obtained using platinum pans with a thin
layer of concentrate spread over the pan surface.  Under these conditions
the sulfur formed was easily removed from the reaction zone by the flowing

-------
TOO
          ATMOSPHERE - PURIFIED  HELIUM
          SAMPLES PREDRIED AT 1 10"C
          100     200      300    400     500     600     700     800     900     1000

                                      SAMPLE TEMPERATURE,  C
        FIGURE 7.   Effect of Heating Rate on  the  Thermal  Decomposition
                    of  Morenci  Concentrate

-------
   100
    96
5   92
    88
    84
    80
55   76
    72
 I
                                              1000C
                                               I
I
I
              100     200     300
400     500    700     700




   REACTION TIME, MINUTES
              800     900
                       1000
               FIGURE 8.   Effect of Temperature on Roasting of Morenci Concentrate

-------
                                   -45-
helium.  When the bulk samples of concentrate were neutral-roasted  the
situation was different.   In the horizontal  tube reactor the depth  of con-
centrate in the tube was  up to 1.5 inches and the bed  was 8-12 inches long.
The argon purge gas tended to flow over rather than through  the bed.  Under
these conditions sulfur removal from the reaction zone was poor.   The same
situation existed in the  pot reactor where the depth of concentrate was
several inches.  Because  removal of the sulfur was so  poor with the bulk
samples they had to be heated for many hours to achieve the same weight
loss that was observed on the thermobalance in an hour or less.
     Each of the concentrates was analyzed by differential  thermal  analysis.
Thermograms obtained with the Morenci, Pima, and Anaconda concentrates  in
argon are shown in Figure 9.  In each case the endothermic reactions up to
about 350C correspond to loss of water from the sample.  The endothermic
reactions between 400 and 700C correspond to the decomposition reactions
which evolve sulfur (and  H^S and SO^).  Above 700C the major endothermic
reactions probably correspond to solid phase changes and/or sample melting.
It is fairly certain that the endotherms at 950-975C  correspond to melt-
ing.  The exotherms which occur on cooling at 950-900C result from freez-
ing:  while in the case of the Anaconda the exotherm at 740C on cooling
is probably a solid phase transition.  The thermograms for the other con-
centrates are similar to  those obtained with the Morenci, Pima, and
Anaconda concentrates.
     A better understanding of the reactions involved  can be obtained by
comparing the DTA and TGA data for a given concentrate.  DTA and TGA plots
obtained with Morenci concentrate in purified helium at a heating rate  of
5C per minute are shown  in Figure 10.  The initial endotherms (up to 350C)
correspond to the initial weight loss due to vaporization of water.  The
large endothermic reaction between 400 and 700C corresponds to the maximum
sample weight loss (evolution of sulfur).  The endothermic reactions above
700C are reversible and  indicate a phase change of one or more constit-
uents of the concentrate.  Some weight loss occurs above 700C but this
would not account for the endothermic reactions observed.

-------
                              -46-
o
o
                       MORENCI  CONCENTRATE
       HEAT-
o
o
                        PIMA CONCENTRATE
        HEAT-
o
x
                     ANACONDA  CONCENTRATE
          HEAT-
        HCATING RATE - 30 C/mln


       J	I	I      I
            -L	I	I      I      I
            200
4CO
600
                    SAMPLE TEMPERATURE, C
800
1000
  FIGURE 9.  Differential  Thermal  Analysis Curves  for Copper

             Concentrates  Heated in Purified Helium

-------
                                        -47-
o
x
LU
O
o
                               DTA CURVES
                  HEAT-
   100
S   95

UJ
X

_l
<


I   90



u.
o

^-
z
UJ

   85
 
-------
                                   -48-
Trace Impurities
     The copper concentrates contain a number of trace impurities which
can play a significant part in a commercial process.   Some of the concen-
trates contain significant precious metal  values (Anaconda and Battle
Mountain).  Others contain impurities which can represent a safety problem
and their fate during the various process operations  must be known.  The
Anaconda concentrate is one which contains significant levels of impuri-
ties; it was used to determine the fate of impurities during neutral
roasting.  The concentrate was neutral-roasted for 24 hours in flowing
argon.  The product was then analyzed for impurities.  The results are
shown in Table VIII, together with the analyses of the original  concentrate.
To account for the weight loss encountered during neutral-roasting the
impurities are reported as grams per 100 grams of original concentrate.
         TABLE  VIII.   Effect  of Neutral  Roasting  at 800C on the
                      Impurities  in Anaconda  Concentrate
                          Grams  per  100 Grams of Original Concentrate
    Impurity
      As
      Sb
      Zn
      Pb
      Mo
      Bi
      Cd
      Se
      Te
      Hg
      Ag*
      Au*
Concentrate
1.86
0.037
6.2
0.05
0.05
0.03
0.12
0.029
0.025
0.024
10.42
0.07
Neutral Roasted
Concentrate
0.17
0.06
3.76
0.2
0.064
0.032
0.50
0.018
0.0056
0.00016
10.2
0.051
* In troy ounces per ton of original  concentrate.

-------
                                   -49-
     From the results it can be seen that the bulk of the arsenic in the
concentrate is volatilized during roasting and should end up in the sulfur.
A fraction of the zinc also appears to have volatilized, probably as the
sulfide.  Conflicting results were obtained with some of the trace level
impurities.  With some elements (i.e., Sb, Pb, Cd) the analytical results
show a greater amount in the neutral-roasted concentrate than in the feed,
clearly an impossibility.  This is due to lack of precision of the ana-
lytical method at very low concentrations.  All that can be said is that
the neutral-roasted concentrate contains a certain amount of the impurity
but the exact level cannot be defined.  In the case of gold and silver the
analyses agree, within the procedure precision, and essentially all of the
gold and silver in the concentrate end up in the neutral-roasted concen-
trate.  This is a desirable situation because it assures that the precious
metals will be recovered from the concentrate by the usual techniques in
the subsequent conventional processing steps of converting and electrolytic
refining.
Sulfur Dioxide Formation
     When the copper concentrate is neutral roasted, some HpS and S(L are
formed.  When the gas stream from the reactor cools the ^S and SCL combine
to form elemental sulfur.  However, the off-gas always contains more S02
than is required to react with the H^S; thus cold off-gas always contains
some SOp.
     The exact mechanisms by which the S(L and H-S are formed have not been
determined, but they probably result from the reaction of water in the con-
centrate with the concentrate and/or elemental sulfur.  This hypothesis is
strengthened by the fact that reducing the water content of the concentrate
prior to neutral roasting reduces the amount of S02 formed during roasting.
     No attempt was made to determine the amount of S02 and H2S formed
during neutral-roasting, but the excess SCL in the off-gas was measured for
Morenci concentrate.  When the concentrate is neutral-roasted at 800C in
flowing helium, approximately 1.2% of the sulfur in the concentrate shows

-------
                                   -SO-
up as free SCL.   The SCL yield was approximately the same when the concen-
trate was neutral-roasted at 1000C.   The concentrate used in these tests
contained about 3.5% water.   When the concentrate was dried under vacuum
at 170C, the water content was reduced to about 1.5%.   When this material
was neutral-roasted at 800C the free S(L formed amounted to about 0.8% of
the sulfur in the concentrate.
     To determine the effect of a reducing atmosphere on the decomposition
reaction and SO,, formation,  Morenci concentrate was roasted in an atmosphere
of Argon-4% Hydrogen at 800C.  Use of a reducing atmosphere increased the
excess SO^ in the off-gas to about 2.1% of the sulfur in the concentrate.
The composition of the neutral-roasted concentrate was  not changed signifi-
cantly from that obtained in argon.
     Two samples of Morenci  concentrate were also roasted at 800C in
He-2% O^  The amount of S0 formed was substantially increased, and the
final products were magnetic.  The composition of the two products were:
                        Sample No. 1          Sample No. 2
          Fe           31.8 wt%             32.86 wt%
         . Cu           24.2                 24.6
          S            26.4                 26.7
                       Fe1.38Cu0.93S2       Fe1.42Cu0.93S2
                       M1.15S               M1.17S
The sulfur content of the samples is lower than usual and it is probable
that some oxidation of the samples occurred, possibly with some sulfate
formation.  The H?S evolved when these materials were treated with HC1 was
also greatly reduced.  This is convincing evidence that the iron sulfide
had been oxidized.
     When pyrite was neutral-roasted, about 1.5% of the sulfur in the
                                                   (79)
pyrite was evolved as SOp.  Work reported by othersv   ' indicated much
higher S0? formation as well as some COS formation when pyrite was neutral-
roasted.  Reasons for the discrepancies between the studies have not been
developed.

-------
                                   -51-
ACID LEACHING
     The leaching of neutral roasted concentrates in hydrochloric acid was
studied in considerable detail using a batch reaction system.   The princi-
pal reaction involved is the dissolution of iron sulfide to form ferrous
chloride and hydrogen sulfide.  The amount of ferric chloride  formed was
usually very low.  The effects of the following variables on the reaction
were evaluated:
 1.  concentrate type (i.e., Morenci, Pima, etc.)
 2.  concentrate composition variations (due to temperature and time for
     roasting)
 3.  reaction temperature
 4.  acid concentration
 5.  acid-concentrate ratio
Two variables which were not studied were stirring rate and concentrate
surface area.  An attempt was made to use the same stirring rate for each
experiment, but the variations in H2S yield and iron dissolution observed
in duplicate runs may reflect variations in stirring rate.   Because the
reaction .involved is a liquid-solid reaction, stirring would have less
effect on the kinetics than if a gas-solid reaction were occurring as in
the case of oxidative leaching reactions involving the use  of  oxygen.  No
attempt was made to study the effect of concentrate surface area on the
leaching reaction.  The only control used was to insure the neutral roasted
concentrates used in the leach tests were all -70 mesh in particle size.
     The progress of a given leaching experiment was followed  by measuring
the HLS evolved.  In some'runs the leach solution was sampled  periodically
and analyzed for iron and copper.  Figure 11 shows typical  data obtained
with neutral-roasted Morenci concentrate and a reaction temperature of
106C (boiling point of 4f[ hydrochloric acid).  By analyzing the feed
material, leach solution, solid residue and off-gas absorber solution
material balances for iron, copper and sulfur were calculated  for the
leaching reaction.  Material balance data for a typical leaching experi-
ment are shown in Table IX.

-------
                                  -52-
    100
o
Ul

s
1

z
^H



5!
tt
OL

=>
     60
     50
     40
     30
     20
      10
                                   IRON
                                 SULFUR
NEUTRAL ROASTED MORENCI CONCENTRATE
            r.

4N HYDROCHLORIC ACID

ARGON COVER GAS


REACTION TEMPERATURE~ 106C
                20
                        COPPER



                            ^-


     40      60       80      100



       REACTION TIME,  MINUTES
                                                       120      140
 FIGURE  11.   Typical  Leaching  Data for  Neutral  Roasted Morenci  Concentrate

-------
                                -53-
TABLE IX.  Material  Balance Data for a Typical  Leaching Experiment
           Using Neutral  Roasted Morenci  Concentrate
Reaction Conditions:
        50.0 grams neutral  roasted concentrate
        300 ml  4j^ hydrochloric acid
        Reaction temperature - 80C
        Reaction time - 150 minutes
Feed Analysis:
        Fe = 33.36% (16.68 grams)
        Cu = 24.66% (12.33 grams)
        S  = 28.8%  (14.40 grams)
Residue Analysis:
        Amount of residue = 21.6 grams
        Fe =  0.58% ( 0.13 grams)
        Cu = 56.4%  (12.18 grams)
        S  = 20.03% ( 4.33 grams)
Solution Analysis:  (Leach solution plus wash = 1000 ml)
      '  Fe = 16.4  g/1 (16,4 grams)
        Cu =  0.46 g/1 ( 0.46 grams)
Scrubber Solution Analysis:
        S = 0.311 moles/1 (9.97 grams)
Material Balances:
        Feed = Residue + Solution + Scrubber Solution
        Fe:  16.68 g = 0.13 + 16.4 = 16.53 g
        Cu:  12.33 g = 12.18 + 0.46 = 12.64 g
        S:   14.40 g = 4.33 + 9.97 = 14.30 g

-------
                                   -54-
     After a few experiments it became  evident  that  the  progress of a
leaching experiment could be followed simply by measuring  the  H,,S evolved.
When H?S evolution stopped or was very  slow, the dissolution of iron was
complete.  The dissolution of copper continued, but  at a slow  rate com-
pared to iron dissolution.  Copper dissolution  did not contribute signifi-
cantly to the generation of HpS.   The rate of H2$ evolution was, therefore,
a good measure of the rate of iron dissolution; it was not necessary to
analyze the leach solution during the course of the  experiment.   It was
only necessary to analyze the final  leach solution for iron and copper
(for material balance purposes).
     In most of the leaching experiments, argon or helium  was  used as  the
purge gas.  It was felt that this would prevent the  oxidation, by air, of
iron (II) in the leach solutions to iron (III).  If  iron (III) is present
in the solution, it could increase the  copper dissolution. The analytical
data showed that in most runs the iron  in solution was present as iron (II).
An exact measure of the iron (III) was  difficult because the  iron  (II) and
total iron usually agreed within the precision of the analytical  procedure.
In some runs air was used as the cover  gas.  In these runs the iron  (II)
and total iron still were the same (within the analytical  precision) and
no increase in iron (III) was observed.  In addition, no significant
increase in copper dissolution was observed.  Therefore, use  of an  inert
cover gas during leaching does not appear to be necessary.
Concentrate Type
     To determine how different concentrates would respond to  acid  leach-
ing, the eight neutral-roasted concentrates were leached under identical
conditions.  The concentrates used had  all been neutral-roasted in  an  inert
atmosphere at 800C for 16-24 hours.
     The conditions used  in the tests were not the optimum for leaching.
Time did not permit determining the optimum leach condition  for each
neutral-roasted concentrate.   Instead,  it was necessary  to select a stan-
dard set of  (less than optimum) leach conditions for use in  comparing  the
eight concentrates.

-------
                                  -55-
     The procedure used was as follows.   Fifty grams  of neutral-roasted
concentrate were placed in the reaction  vessel  and  the vessel  purged  with
argon.  Three hundred ml of 4N^ hydrochloric acid (M00% excess)  were
added, and the system heated to the boiling point of  the acid  (^106C).
It took approximately 10 minutes for the acid to reach the boiling  point.
Each run lasted 90 minutes.  H2$ evolution was essentially complete after
90 minutes with each concentrate.
     The results obtained with the various concentrates are given in
Table X.  It is readily apparent from the data that since at least  67% of
the sulfur in the original concentrate needs to be  converted to H,,S that
none of the neutral-roasted copper concentrates produced, under the leach-
ing conditions used, the amount of H2S needed for a workable process.  The
copper concentrates which are high in pyrite (Morenci and Tyrone) come the
closest to meeting process requirements.  The San Manuel, Pima and  Anaconda
concentrates, which are low in pyrite, produce the  least HgS.   The  Hecla
concentrate (which is principally pyrite) after neutral-roasting reacted
almost completely with the acid.  The residue remaining after leaching con-
tained only trace levels of iron and sulfur.
      From the results presented, it is apparent that in order to obtain a
workable process more H^S must be produced from the copper concentrates.
There are several ways in which this might be done.
 1.   Optimizing the leaching conditions.
 2.   Varying the neutral roasting conditions to change the composition of
      the neutral-roasted concentrates.
 3.   Add neutral-roasted pyrite to the neutral-roasted concentrate to
      increase the H2S formation.
All  three approaches were  investigated using neutral-roasted Morenci, Pima
and  Anaconda concentrates.
      In the laboratory  studies  the leach  solution and  solid residue were
separated  by filtrations  using  a #42 paper.   In a commercial operation
filtration would  probably  also  be  used for  the  separation.  In  the plant

-------
                                    -56-
      TABLE X.  Leaching Data for Neutral Roasted Concentrates
          Reaction Conditions:
                    . .^T-'-i
               Temperature

               Reaction Time

               Concentrate

               Hydrochloric Acid -
:Wl06C
  90 minutes

  50.0 grams

  300 ml  4N  acid
Neutral Roasted
Concentrate*
Morenci
Pi ma
Anaconda
Tyrone
Lavender Pit
Battle Mountain
San Manuel
Heel a**
Amount of Material in Concentrate
Reacted, %***
Fe
87.7
62.5
74.9
83.4
65.0
85.3
64.6
100.0
Cu
1.80
1.90
1.10
2.90
2.70
4.10
4.13

S '
60.6
41.0
47.9
61.2
52.3
57.1
40.1
100.0
  * All concentrates neutral roasted at 800C for 16-24 hours in
    helium or argon.  See Table II for analyses.

 ** Only 25.0 grams of neutral  'roasted concentrate used.
***
    Average of two or more runs: iron and copper dissolved in
    leach solution and sulfur evolved as HS.

-------
                                   -57-
the residue would have to be thoroughly washed to remove  chloride  effec-
tively before further processing to recover the copper.   The filtration
characteristics of the leach residue are,  therefore,  an  important  factor
to be considered.  It was found that, with the exception  of the neutral -
roasted Pima, the leach residues were easily filtered and washed.   Con-
flicting results were obtained with the Pima concentrate.  In most runs  the
residue was as easy to filter and wash as the other concentrates.   In a  few
runs, however, the residue was extremely difficult to filter.  The cause of
the filtration difficulties has not been identified.
Concentrate Composition
     The composition of a neutral-roasted concentrate varies depending on
the conditions of roasting.  This variation in composition will in turn
affect the results obtained during acid leaching.  To determing how composi-
tion would affect the H2$ yield, samples of neutral-roasted Morenci, pre-
pared under various conditions, were leached  under identical conditions.
The leach procedure was identical to that described in the previous section
(fifty grams of  neutral-roasted concentrate were leached in 300 ml of boil-
ing 4N^ HC1 for 90 minutes).
     The results obtained with  neutral  roasted Morenci concentrate are
given in Table XI.  The results show that the H2S yield  (and iron dissolu-
tion) is increased by increasing the metal-sulfur ratio  of  the neutral-
roasted feed.  This ratio  is best increased by increasing the  neutral-
roasting temperature, lengthening the  roasting time, or  both.  Maximum h^S
yield was obtained when the  neutral-roasting  temperature was high enough  to
melt  the concentrate  (^1000C).
      Differences in  the mineralogical  composition of concentrate  roasted  at
different  temperatures may also help explain  the differences  in the  H2S
yield.   Since  the mineralogical  composition of the neutral-roasted concen-
trate could  not  be adequately  determined,  however, it was  impossible  to
evaluate  this  effect.

-------
                                    -58-
         TABLE XI.  Effect of Concentrate Composition on Leaching
            Leaching Condition:
               Concentrate   - 50 g neutral roasted Morenci
               Acid          - 300 ml 4N_ HC1
               Temperature   - ^ 106C
               Reaction Time - 90 minutes
Sample
No.
1
2
3
4
Roastim
Temperati
800C
800C
970C
1005C
Results
Sample
No.
1
2
3
4
3 Composition, wt%
jre Fe Cu S
32.3 23.7 29.2
33.4 24.7 28.8
37.3 28.5 30.5
33.5 25.9 25.2
of Leaching Tests:
Amount of Material in
Concentrate Reacted, %*
Fe Cu S
75.9 2.9 52.3
87.7 1.8 60.6
88.6 0.97 62.2
95.6 4.0 71.1
M/S
M1.05S
M1.10S
M1.17S
M1.23S
* Fe and Cu dissolved in leach solution, S evolved as H?S.

     It was found that Morenci concentrate neutral-roasted at 1005C was
much more reactive than that produced at lower temperatures.   Data presented
in Figure 12 shows that the concentrate prepared at 1005C reacts more
rapidly with acid at room temperature than material  prepared  at 800C does
at 60 an.d 106C.  Another surprising factor is that the temperature of
leaching has only a slight effect on the reaction rate for the material
prepared at 1005C.

-------
                               	MORENCI  ROASTED  AT  80CTC
                                       MORENCI  ROASTED  AT  1005"C
                               50g CONCENTRATE


                               300 ml  4N  HCl
                       40      50      60      70


                       REACTION TIME, MINUTES
                                                                      100
i
en
FIGURE 12.   Reactivity  of  Neutral-Roasted Morenci Concentrate

-------
                                    -60-
     Neutral-roasted Pima and Anaconda concentrates prepared at various
temperatures were also leached to determine the effect of composition on
H2$ generation and iron dissolution.  Leaching conditions were identical
for each experiment and were the same as those used with the Morenci con-
centrate.  The results obtained are presented in Table XII.  Again it was
found that, in general, the higher the metal-sulfur ratio in the neutral-
roasted concentrate (of a given type) the higher the HLS yield and iron
dissolution.  Increased roasting temperature also increased the reactivity.
     Copper dissolution during leaching of the various concentrates was
erratic.  In every run in which periodic samples were taken copper dissolu-
tion increased with time.  However, there was considerable variation in the
amount of copper dissolved in similar or duplicate runs.  Reasons for the
variability in copper dissolution have not been determined.
     Tests with neutral-roasted Heel a pyrite showed the material  dissolved
almost completely in hydrochloric acid with essentially complete dissolution
of the iron and conversion of sulfur to H^S.  One way of increasing the HpS
yield from neutral-roasted copper concentrates would be to leach a mixture
of neutral-roasted pyrite and neutral-roasted copper concentrate (pyrite is
usually available as a waste stream from the flotation plant in most copper
operations).  Roasting the pyrite and concentrate separately would insure
sufficient H^S production.  A second possibility would be to combine the
pyrite and concentrate prior to roasting.  To see if this procedure could
be used, mixtures of Heel a pyrite with Morenci, Pima and Anaconda concen-
trate were neutral-roasted in helium at 820-860C for 16 hours and then
leached.  The results obtained are presented in Table XIII.  The addition
of pyrite to the Pima concentrate increases the H^S yield on leaching to an
acceptable level.  With the Anaconda concentrate there was a small  but
significant increase in the hLS yield.  In the case of the Morenci, the H?S
yield decreased slightly but was within the reproducibility of the experi-
mental  procedure.  One factor noted was that the reaction rate for neutral
roasted concentrate-pyrite mixtures decreased compared to straight neutral

-------
                                     -61-
      TABLE  XII.   Effect  of Composition on  Leaching of  Neutral
                    Roasted Pima  and Anaconda  Concentrates

                 Leaching Conditions:
                   50.0 g neutral  roasted concentrate
                   300 ml 4J1 HC1
                   Temperature -v. 106C
                   Reaction  Tine 90 minutes
Concentrate
Pima
P1ma* A
B
Pima
P1ma
Anaconda
Anaconda
Roasting
Temp.
C
800
860
860
970
1005
800
1000
Comp. of Neutral Roasted Cone.
w
Fe
29.4
29.1
28.9
28.9
29.1
23.2
28.1
wt%
Cu
29.4
30.9
28.6
29.7
29.3
34.9
42.3
wt%
L S
27.8
25.5
24.6
24.7
24.8
26.5
26.9
M/S
M1.14S
H1.25S
M1.27S
M1.28S
M1.27S
M1.17S
M1.39S
Amount of Material
in Cone. Reacted ,_*,*
fe
62.5
92.2
86.3
88.2
92.4
74.9
85.6
Cu S
1.90
1.3
0.6
0.6
1.6
1.1
4.8
41.0
64.8
56.3
59.6
59.7
47.9
51.8
     * Fe and Cu dissolved in leach solution and S evolved as  H?S.

     ,   TABLE  XIII.   Leaching of  Neutral  Roasted Pyrite-Copper
                       Concentrate  Mixtures
    Leaching CunuiL'iu.is:
       50.0 g neutral roasted concentrate
       400 ml 4H HC1
       Temperature ^  106C
       Reaction time up to 120 minutes

    Leaching Results:


Feed to Roaster
80Z Morenc1-20 Pyrite
57. 5X Pima-42.5% Pyrite
50% Anaconda-SOS Pyrite
Roasting
Temp.
(C)
860
820
860
Roasting
Time
(hrs)
16
16
16
Comp. of Neutral
Roasted Cone. ,
Fe
35.4
39.5
35.0
Cu
21.1
30.0
20.5
wtS
S
28.2
28.8
28.0
Amount Reacted 1n
Leaching, %*
Fe
84.2
79.1
71.4
Cu
1.1
2.5
2.8
S
59.3
68.0
54.1
* Fe and Cu dissolved in leach solution and S evolved as H2S.

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                                     -62-
roasted concentrates.  Normally at 106C the leaching reaction was essen-
tially complete in 60 minutes or less.  With the neutral-roasted mixtures
it required 90-120 minutes for completion.
     Based on the results of the leach tests with concentrate-pyrite mix-
tures, it appears that separate roasting of the concentrates and pyrite
would be preferred to combined roasting.
     Samples of Morenci concentrate which had been roasted at 800C in
He-2% 02 were leached in 4N. HC1 at 106C.  The H2S yield was greatly
decreased as was the iron dissolution.  The maximum KLS yield obtained was
only 35%.  In addition the reaction rate was greatly decreased.  It
required approximately 10 hours before the evolution of H2S stopped.  Iron
dissolution was less than 60% and copper dissolution less than 1%.  These
results are difficult to explain.  If oxidation of the samples occurred
during roasting, the iron dissolution should be much higher.  Similarly if
sulfation occurred during roasting, both iron and copper dissolution should
be higher.  The only conclusion that can be made is that the presence of
oxygen in the cover gas during roasting can adversely affect the process by
reducing the HpS yield obtained on roasting.
     When samples roasted in Ar - 4% H2 were leached, the yields obtained
were similar to those obtained with concentrate roasted in argon.   There-
fore the use of a reducing atmosphere during roasting does not appear to
adversely affect the process.
Operating Variables
     The principal  operating variables which affect the leaching reaction
are:
 1.   acid concentration,
 2.   acid-concentrate ratio,
 3.   reaction temperature.
The  effect of each variable on the leaching of roasted Morenci, Pima, and
Anaconda concentrates was studied.   The concentrates used were all  neutral
roasted at 800C in flowing argon for 24 hours (see Table II for analyses).

-------
                                    -63-
     It was found that the acid concentration used had a significant effect
on the leaching of neutral roasted Morenci concentrate (the maximum acid
concentration tested was that of the azeotrope 'vSM).   Maximum iron dissolu-
tion and h^S formation was obtained with an initial  acid concentration of
3-4 molar (see Figure 13).  A 100% excess of acid (based on iron content
of the feed) was used in each run.  Both iron dissolution and H?S produc-
tion dropped sharply as the initial acid concentration was varied from the
optimum.  The dissolution of copper increased rapidly with increasing acid
concentration and was especially severe when the initial acid concentration
was greater than 4M.  The amount of undissolved solid residue from the
leaching reaction varied with acid concentration and  was a minimum, as
expected, when H^S production and iron dissolution were a maximum (see
Figure 14).
     The initial  acid concentration has far less effect on the leaching of
neutral-roasted Pima concentrate as far as H^S production and iron dissolu-
tion is concerned (see Figure 15).  Both H?S formation and iron dissolution
increased slightly with increased acid concentration.  As was the case with
Morenci concentrate, copper dissolution increased rapidly with increasing
acid concentration.
     With neutral-roasted Anaconda concentrate maximum iron dissolution and
HpS production were obtained with an initial acid concentration of about 4M^
(see Figure 16).   Changes from the optimum initial acid concentration caused
only a slight decrease in H^S formation.  Copper dissolution increased with
increasing acid concentration but not to the same degree as with Morenci
and Pima concentrates.
     The ratio of acid to concentrate used can affect the H^S yield obtained,
With neutral-roasted Morenci, a 100% excess of acid  (in excess of that
needed to react with the iron in the feed) was required to obtain the maxi-
mum HpS yield and iron dissolution (see Figure 17).   With neutral  roasted
Pima, H2S formation and iron dissolution were relatively independent of the
excess acid used (see Figure 18).  The neutral-roasted Anaconda concentrate
gave a slight increase in HpS production with increasing acid-concentrate
ratio.

-------
                               -64-
 o
 QC
    100
     90
     80
     70
     60
     50
    40
<
LU
oc
o
o

-------
                                   -65-
     75
oc.
UJ
D-

 ft
LU

O
li
LO
UJ
o:

:r
o
O

to
HH
o
     70
     65
     60
     55
     50
     45
     40
                        INITIAL ACID CONCENTRATION, M
      FIGURE 14.  Effect of Initial Acid Concentration on Dissolution
                 of Neutral-Roasted Morenci Concentrate

-------
                                -66-
CJ

Of
o

-------
                                -67-
o
OL
Q.


 >

O
o
o
o
z:
ct
    80
    70
    60
    50
    40
    30
    20
    10
TEMPERATURE- 106C
TIME - 90 MINUTES
25g CONCENTRATF

.1002 EXCESS HC1
                              COPPER
       123456


                      INITIAL ACID CONCENTRATE,  M



  FIGURE  16.   Effect of  Initial Acid Concentration on the Leaching
              of Neutral-Roasted Anaconda Concentrate

-------
                                -68-
a:
UJ
a.

-------
                                   -69-
 o
 ce
a
ui
UI

Of
o
o
      50
      40
      30
E    20
     10
                              IRON
              TEMPERATURE  ioec
              TIME - 90 MINUTES
               25
50
                               75
                   SULFUR
                                        O
                                          'COPPER
               100
                                              125
                               150
                                                              175
                        fXCESS  ACID  USED,  PERCENT
   FIGURE 18.  Effect of Excess Acid on Leaching of Neutral-Roasted
               Pima Concentrate

-------
                                     -70-
      The reaction temperature has a significant effect on the leaching
 reaction.  The effect observed was unexpected in that maximum hLS production
 and iron dissolution were obtained at temperatures below the boiling point
 of the acid solution.  The results obtained with the three concentrates  are
 shown in Figures 19-21.   The tests were carried out by bringing the 4f1 HC1
 to the reaction temperature and then adding the concentrate.   The runs were
 continued until  the evolution of H2$ was essentially complete.   Maximum  H?S
 production was obtained  with each concentrate when the reaction temperature
 was between 80-90C.
      With neutral  roasted Morenci  the rate  of H2$ formation  increases  with
 increasing reaction temperature as shown in Figure 22.   At a  temperature of
 106C,  H2S formation  and iron dissolution is  complete  in  30 minutes  or
 less.   At 70C and  below the reaction rate  decreases rapidly  and  reaction
 times  of 4 hours  or longer  are required.  At  80C,  where  maximum  H  S forma-
 tion  is  obtained  (with Morenci),  the  reaction  is  complete  in  about 60-70
 minutes.   The  reactivity of neutral-roasted Anaconda is similar to that of
 the neutral-roasted Morenci,  while  reactivity  of  the neutral-roasted Pima
 was somewhat less.  With  each  concentrate maximum  H_S yield is obtained at
 a  reduced  reaction  temperature  at  the expense  of  increased reaction  time.
     One  additional factor  was  noted with regard  to reaction  temperature.
 If the acid and concentrate were combined and  heated to the reaction tem-
 perature,  the  H2S yield was higher than when the concentrate was added to
 the acid  at the reaction  temperature.  This was especially true at reaction
 temperatures of 90C and above.  The reason for this phenomenon has not
 been resolved.
 Fate of  Impurities During Leaching
     Samples of solution and residue from a  leaching experiment with neutral
 roasted Anaconda concentrate were analyzed to determine the fate of impuri-
 ties during leaching.  Table XIV shows how the impurities divide between
 the leach solution and solid residue.  Data  for the original  concentrate
and neutral-roasted concentrate are also shown (to put the data on a uniform

-------
                                -71-
    100
     90
     80
o
oc
o
LLJ
     70
     60
     50
O
O
     30
     20
     10
                                          COPPER
           60
70
80
90
100
110
                        REACTION TEMPERATURE, C
     FIGURE  19.   Effect of Temperature on Leaching of Neutral -
                 Roasted Morenci Concentrate

-------
                                 -72-
CJ
ce.

-------
                                -73-
 o
 cc
O

-------
                                  -74-
I/O
 CM
o:
o   40  -
o:
ID
t/0
                               100% EXCESS  4N  HC1

                               50.Og  NEUTRAL  ROASTED  MORENCI
               50      TOO      150      200      250

                        REACTION TIME,  MINUTES
300      350
        FIGURE 22.   Effect of Reaction Temperature on the Leaching
                   of Neutral-Roasted Morenci Concentrate

-------
                                     -75-
  TABLE XIV.  Fate of Impurities During Leaching of Anaconda Concentrate
Impurity
As
Sb
Zn
Pb
Mo
Bi
Cd
Se
Te
Hg
Ag*
Au*
Grams Per 100 Grams of Ori<
Concentrate
1.86
0.037
6.2
0.05
0.05
0.03
0.12
0.029
0.025
0.024
10.42
0.07
Neutral Roasted
Concentrate
0.17
0.06
3.76
0.20
0.064
0.032
0.50
0.016
0.0051
0.00016
10.2
0.051
inal Concentrate
Leach
Residue
0.074
0.05
0.20
0.20
0.005
0.006
0.0025
0.0094
0.0035
0.0002
2.24
0.06
Leach
Solution
0.084
0.10
3.12
0.32
0,005
0.008
0.0016
0.0064
0.0024
0.000016
7.70
0.012
  *In  troy ounces  per ton  of  original  concentrate.
basis the impurity levels are reported as grams per 100 grams of original
concentrate).  The overall material balances for the trace level impurities
are poor due to analytical problems.  It is apparent, however, that the
trace impurities divide between the leach solution and the solid residue.
The bulk of the zinc ends up in the leach solution as does the silver,
while the bulk of the gold remains in the residue.
     When the leach solution is treated  with H2$ most of the copper pre-
cipitates as does the silver and whatever gold  is present (see Table XV).
The effect on the other impurities in the leach solution is  minimal.

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                                     -76-
                TABLE XV.  Effect H2S Treatment on Impurities
                           in the Leach Solution
Impurity
Cu
As
Sb
Zn
Pb
Mo
Bi
Cd
Se
Te
Hg
Ag
Au
Original
Leach Solution
0.92
0.105
0.125
3.90
0.40
0.006
0.010
0.002
0.008
0.003
0.00002
0.033
0.00005
Treated
Leach Solution
0.05
0.06
0.10
4.20
0.40
0.005
0.002
0.008
0.009
0.002
0.00001
0.002
Trace
Process Optimization
     The laboratory studies have shown that it will  be possible to achieve
sufficient H2S production with most types of copper  concentrates to make
the proposed process viable.   It will  require, however,  optimization of
the neutral-roasting and acid leaching operations and pyrite addition to
insure adequate H2S production from low iron concentrates.
     To maximize H2S production with all  types of concentrates,  the follow-
ing process  conditions  should be met.
 1.  The neutral  roasting operation should  be carried out at the highest
     temperature possible (>800C).  Melting of the  concentrate  during
     roasting  produces  the optimum  product  from a  reactivity and H?S
     generation  standpoint.

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                                     -77-
  2.   Care must be taken to exclude oxygen from the cover gas during roast-
      ing.  A neutral or reducing atmosphere must be used to obtain maximum
      H2$ production upon leaching.
  3.   If possible, a mixture of neutral-roasted copper concentrate and
      neutral-roasted pyrite should be used for leaching.  This will insure
      adequate H2$ production.  Since a typically sized copper smelter would
      require more than one roaster, the copper concentrate and pyrite can
      and should be roasted separately.
 4.   The leaching should be carried out with an initial  acid concentration
      of about 4N.at a temperature of 80-90C.   An excess of acid should be
      used.   The excess required will  depend  on the concentrate but will
      probably be at least 50% and possibly as  much as  100%.
 5.   The residence time of the concentrate in  the leach  tank should be the
      shortest possible time consistent with  the required H2$ production.
      Increasing the  residence time will  increase  copper  dissolution.
If the process  operation conforms to  the  conditions  set  forth  above,  the
H^S production  will  be sufficient for a  viable process.

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                                     -78-
                        ECONOMIC ANALYSIS OF PROCESS

     The cost of utilizing the proposed process for a 300 tons per day copper
smelting operation was estimated based on the data developed in the laboratory
studies.  Two estimates were prepared:  one assumed the use of a chalcocite
(Morenci) concentrate feed and the second a chalcopyrite (Pima) concentrate
feed.  In each case pyrite addition was used to increase the availability
of H^S.  The pyrite concentrate was assumed to be available as a by-product
stream from the flotation mill.  There may be some question concerning the
real need for additional  pyrite for Morenci concentrate processing, however
pyrite addition is required with the Pima concentrate.   It was used in the
Morenci case simply to insure adequate KLS formation.   The pyrite would be
roasted separately from the copper concentrate, and combined with the neutral
roasted copper concentrate prior to leaching.
     The following assumptions were also made in making the cost estimates:
 1.   The copper-bearing residue from the leaching operation would be fed
     directly to a converter.   This eliminates the need for a reverberatory
     furnace.
 2.   HC1  recovery and conversion of ferrous chloride to ferric oxide would
     be carried out in equipment similar to that currently being used to
     process hydrochloric acid pickle liquor  '  in the  steel  industry.
 3.   The reaction of H-S  and  SO- to form elemental  sulfur  would be carried
     out using the "citrate"  process  developed by the U.S.  Bureau of
           (36}
     Mines.       If some  SO^  release  could be  tolerated, a  modified Claus
     process could be used for the reaction at a somewhat  lower cost.
 4.   The precious metals  in the concentrates would end  up  in  the blister
     copper  from the converter.   They would be recovered when the copper
     is refined by electrolysis.
 5.   The  impurities in the concentrate,  other  than the  precious metals,
     were ignored in preparing  the cost  estimates.

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                                     -79-
     Process flow diagrams were prepared for the two cases under considera-
tion and are shown in Figures 23 and 24.  Stream flows were calculated using
the data developed in the laboratory studies.  The bases used in the calcu-
lations are shown in Table XVI.  Impurities in the concentrates were not
considered in the preparation of the flow diagrams.
     For the purpose of the cost estimates, the overall smelter process was
broken down into five major operations.  Capital and operating costs for
each major operation were estimated separately (the various subsidiary
operations were included in the first estimates for the principal  operations),
The five major operations are:
 !  Neutral  Roasting System -  This includes storage,  drying and neutral
     roasting of the concentrates;  particulate removal  from the roaster
     off-gas; condensation, collection and  stockpiling  of the elemental
     sulfur;  and quenching and  granulation  of the neutral  roasted
     concentrates.
 2.  Acid  Leaching  System - This operation  includes blending of the  neutral
     roasted  concentrates; leaching of the  concentrates;  filtration  of the
     leach slurry and washing of the solid  residue; treatment of the leach
     solution with  H,,S to precipitate the dissolved copper  and precious
     metals;  filtration  of the  leach solution  to collect  the precipitated
     copper and  precious metals;  washing of  the  precipitate;  and collection
     of the H^S  from  the leach  circuit.
 3.  Acid  Recovery  System -  This  includes evaporation of  the filtered  leach
     solution to remove  the  excess  water; conversion of the ferrous  chloride
     to ferric oxide  and hydrogen chloride;  collection  of the  hydrogen
     chloride as 4  to 5N^ hydrochloric  acid;  and  handling and  stockpiling
     of the ferric  oxide.
4.   Sulfur Dioxide Collection  and  Sulfur Production -  This  operation
     includes collection of  the  SO^-containing gases from the  converter;
     handling of the  hydrogen sulfide  gas stream; reaction  of  the S0? and
     H2$ in an aqueous citrate  system  to form elemental sulfur;  stockpiling
     of the sulfur; and  stack-venting  of the SOp-bearing off-gas (approxi-
    mately 0.4% of the  sulfur  in the  total  plant feed  ends up in the
     stack  gas from the  SCL  recovery unit).

-------
        TO STACK
                    S : 1986 1/D

                  TO STACK
                  H70 : 83 T/D
MORENCI CONCENTRATE
   1365 TONS/DAY
Fe
Cu
S 
 26%  354.9 T/D
  212%  303.3 T/D
35%  477.8 T/D
H20  8*  109.2 T/D
                                                  S2 = 24.1 TH
                                                Slas Sty : a9 T/D
                                          S2 i 174.5 T/0
                                          SOs SO?). 7.2 T/D
                                               1074.1
                                                                      TO STACK"
                                                                                           137 T/D
              NATURAL CAS .
              5.430 MSCF 10
                                             1175.1 T/D
                                             Fe =  404.2 T/D
                                             Cu . 303.3  T/D
                                             S = 328.6 T/D
                                  H,S
                                                        kATURAL
                                                          CAS	1
                                                       410MSCF/D
PYRITE CONCENTRATE
   137 TONS/DAY
Fe  36%   49.3 T/D
S  42%  57.5 T/D
H?0   8%  11.0 T/0
                                                                         -MAKEUPHCI
                                                                                            TO STACK
TO STACK
SOS
          2.2 T/D
   SULFUR

 S- 325.8 T/D
     TO STACK '
     SlasSC^I. CL6T/D

  COPPER ANODES'^	
                             Slas H2SI. 218 T/D

                                    H2S
                                                                    4N HCI
                                                       CIRCUIT'!*     .2    i
                                                       T	'   1.63x106 CAL/D
                           Slas H2S>= as T/D

                    SOs H2SI. 217.2 T/D
                                                                                          Fe2

                                                                                           372.4 T/0
                                                                                       Cu . 0.1 T/D
                                                                                                 2.1x105 Cal/D
                                                                                                 1.87x106 CAL/0
                                                                                                 1.79N HCI
                                                                                                     FeCl2
                                                     FILTER
                                            LEACH RESIDUE
                      StasSO?)  110.8 T/D
                                            Cu - 300 T/D
                                            Fe = 31.8 T/0
                                            S : 110.6 T/D
[
R

WASH WATER
2x.05 GAL/D
1.82X106 GAL/D
Fe = 372.4 T/l>
1


COPPER
RECOVERY
NATURAS
MSCf-/
WASH
SxlO4

CAS
D
WATER
CAL/D if

                                                                                                        3.2 T/D
                                                                                                        as T/o
                                                                                SLAG
  Cu i 300 T/D

t
HOLDING
FURNACE


i
t
FLUX
50 T/
                                                                          *cIT
                                                                            Fe.
                                                                                   . 3.2 T/0
                                                                                   31.8 T/0
                                                                                                                                      oo
                                                                                                                                      o
                                                                                                                                       i
     FIGURE 23.   Process  Flow  Diagram for  300  T/D Copper  Smelter Using Morenci
                     Concentrate Plus  Pyrite Concentrate

-------
TO STACK (
Slas S02*z7.4
, 	 SULF
10 CONOEr
SULFUR
S i 141.7 T/D
 TO STACK

PIMA CONCENTRATE
122 TONS/DAY
Fe = 25.0% - 300.5 T/D
Cu * 255%  306.1 T/D
S = 28.0% - 336.6 T/D
H20 = 8.0% = 96.1 T/D
H20 = 80 T/D
NATURAL GAS_^
4.530MSCF/D
TO STACK
Slas SC^fc 2.3 T/D
$ 340.8 T/0
TO STACK
SCISSOR. 0.6
COPPER ANODES
Cu = 300 T/D
SI
S(a$H2S
SULFUR
RECOVERY
Slas SO?)

/U *
REP
FUfl
UR
ISER
Sla
-I H22 I NEUTRAL
-1 T/D | ROASTER

t
H2S
as H2S). 228 T/D
H2S


S? = 65.1 T/0 '
Slas SOz) : 2.4 T/0
76.6 T/D
 SO?) . 5.0 T/D
1024.3
T/D
1299 T/D
Fe . 434.4 T/D
S = 343.7 T/D
:u . 306.1 T/D
i

las H2S) = as T/0
) . 227.2 T/0
H20
23.8 T/D

 274.7 NEUTRAL 4 3187 , ' u 372 T/D
*~iyb~RnASTER*lF~lRIER r*-^ 
NATURAL
1.130MSCF/D

r MAKEUP HCI
.TO STACK
H.n. 7 i.
*~m. SNHCI 4 HCI .
1 	   , 	 ' i ru.iM r.Ai in RECOVERY


| FILT
Cu -302.8 T/D
Fe - 57.9 T/0
S i 115.7 T/D
= 115 9 T/D f^^

T T

INING. HOLDING
MACE FURNACE

'
 > I i-e2 03
1 . Fe - 376.5 T/D
Cu . n i i/n
WASH WATER HCI NATURAL CAS
2xl01) GAL/0 MSCF/O
, WASH WATER
Jf-| WLIU. COPPER 5xl(^CAL/D r^-
r ' Fe . 375.6
Cu : 3.3 1
T/D' RECOVERY 'U^
/D
1
imFI k SIAG
:!iHJ *" Cu  6.0 T/D
T F . 57.9 T/D
\ FLUX
90 T/0
PYRiTE CONCENTRATE
372 TONS /DAY
Fe . 36% . 133.9 T/D
S . 4?% . 156.2 T/D
H?0 - 8* : 29.8 T/D
ID^Cal/D
1.23X106 Cal/0;
1.56N. HCI
1.26\1 FCI2
TER~I
Cu . 3.2 T/0
S- 0.8 T/D
                                                                                           I
                                                                                          CO
FIGURE 24.   Process  Flow Diagram for a 300 T/D Copper Smelter
            Using Pima Concentrate Plus Pyrite Concentrate

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                                         -82-
               TABLE XVI.   Bases  Used in  Preparing  Flow Diagrams
                                                        Morencl    Pima
                                                        Smelter   Smelter
                1. Copper Production (tons/day)                 300      300
                2. Feed-Concentrate (tons/day)                 1365     1202
                3. Feed-Pyrite (tons/day                      207      429
                4. Drier Temperature (C)                      200      200
                5. Concentrate Roaster Temperature (C)            900      900
                6. Pyrite Roaster Temperature (C)               800      800
                7. S 1n Concentrate Converted to SO, in             15       15
                   Roaster (%)                i
                8. S In Pyrite Converted to S02 in Roaster (%}       1.6       1.6
                9. Leach Acid Concentration (Nj                  4        5
               10. Excess Acid Used (?)                        88      40
               11. S in Neutral Roasted  Concentrate              62.6     55
                   Converted to H2S (S)
               12. S in Neutral Roasted  Pyrite Converted           100      100
                   to t^S (%)
               13. S02 Conversion to Elemental Sulfur (%)          98      98
               14. S Content of Blister  Copper (wtJ)               0.2      0.2
 5.   Converter Operation - This includes operation of  the converter to  form
      blister  copper,  fire refining of  the blister copper, casting into  anodes,
      handling and disposal of  the converter slag, collection of  the SCL-bearing
      gases, cooling and cleaning of the  gases  in  a cyclone and waste heat
      boiler.
      The products from the process are copper  anodes containing  the precious
metals;  ferric oxide  of fairly high purity; pure  elemental sulfur from  the
S02  recovery  unit; and impure  elemental  sulfur  from the  neutral  roasters.
      If  the elemental  sulfur is to be marketed, the sulfur from  the neutral
roasters would require additional  purification.   Add-on  equipment could  be
used  for the  purification at a  nominal cost.  The ferric oxide should be of
suitable purity for direct marketing since the  residue impurities known  to

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                                        -83-
be  present  in  the original concentrates which may carry over  into the  iron
oxide are of  little significance  in iron oxide intended for use in iron
making.  The  principal  residual elements such as zinc, lead,  arsenic and
antimony are  expected  to be of no consequence in iron  making  which would
be  the intended market for by-product iron  oxide.
      Wherever  possible,  equipment sizing, utility requirements, capital
costs and labor requirements are  based on published  data for  similar opera-
tions.   Where  published  data are  not available, capital  costs were estimated
by  standard cost estimating procedures.  Operating costs were estimated
using the data  presented in Table XVII.  In  general  a  conscious effort was
made  to be conservative  in estimating both  capital and operating costs.
It  was  felt that this was  necessary because  of the lack  of pilot and plant-
scale data for  some of the major  operations.   With additional data, it
would probably  be possible to reduce both capital and  operating costs
significantly.

           TABLE  XVII.   Basis for  Estimating  Plant Operating Costs
                         (330 day per year operation)

                     1. Direct Labor (Including fringe benefits)
                       A. Operating  . $5.00/hr
                       8. Maintenance - 31 of fixed capital costs
                       C. Supervision - 181 of operating and maintenance labor
                     t. General Plant Overhead - 2/3 of direct labor
                     3. Utilities
                       A. Power      -  $0.01/h*
                       B. Natural Gas  -  $0.75/1000 ft3
                       C. Cooling Water -  $0.03/1000 gal
                       0. Process Water -  $0.20/1000 gal
                       E.  Boiler Water -  $0.75/1000 gal
                       F.  Steam      .  $0.50/1000 pounds
                    4. Maintenance Supplies and Parts - 1.5S of fixed capital  costs
                    5. Operating Supplies - 10S of direct labor
                    . Pyrite - $Z/ton
                    7. Taxes and Insurance - 21 of plant cost
                    B. Depreciation - 15 year-straight line

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                                    -84-
 NEUTRAL ROASTING SYSTEM
      Storage and handling  facilities  would  be  required  for the  copper  con-
 centrate and the pyrite concentrate.   Although drying may  not be  needed,  it
 is included as  a step  in the  process  in  order  to  minimize-reactions  which
 may increase the sulfur dioxide  content  of  the neutral  roasted  flue  gas.
 Drying  of the concentrates would be accomplished  in  separate direct  fired
 fluid bed driers.   Each drier would be fired with natural  gas.  Maximum
 drier temperature would be 200C.  The off-gas from  each drier  would pass
 through a cyclone for  particulate  removal and  then be vented to the  stack.
 The dried concentrates  would  be  fed to gas-fired  fluid-bed  roasters.   The
 pyrite  roaster  would operate  at  800C.   Residence time  in  the roaster  would
 be 30 minutes.   Two copper concentrate roasters would be used,  each  operat-
 ing at  900C.   Residence time in  the  roasters  would  be  one  hour.  The
 reactors  would  be fired with  a rich fuel mixture  to  reduce  the  oxygen  in
 the product  gas.  The off-gas and  elemental sulfur from the roasters would
 pass  through  cyclones,  for particulate removal, and  then be combined.  The
 combined  gas  stream would pass through a waste heat  boiler, electrostatic
 precipitator  and into a sulfur condenser.  The liquid sulfur would be  col-
 lected  and granulated.  A waste  heat  boiler for low  pressure steam would
 be  operated  in  conjunction with  the sulfur condenser.  The clean off-gas
 would be  vented to the stack.   The SO,, in the off-gas would be approximately
 1.5%  of the sulfur in the combined feed.   The neutral roasted products would
 pass  from the roasters into a quench  tank where they would be cooled and
 granulated (if  necessary).   Quenching could be in water.
     Capital and operating  costs for  the neutral roasting operation are
 presented in Tables  XVIII and XIX.  Equipment,  labor and utility require-
ments are based primarily on a paper published  in 1968 by Mehta  and Okane^28'
entitled  "Economies  of Iron and Sulfur Recovery from Pyrite."  In  the esti-
mates, allowances were made for inflation since 1968, higher operating
temperatures of the  roasters,  and duplicate equipment requirements.   Operat-
ing costs were calculated using the figures presented in Table  XVII.

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                                   -85-

       TABLE XVIII.  Capital Cost for Neutral Roasting Operation

                                                        Cost SlOOO's
                                                     Morenci      Pima
 1.  Purchased Equipment Cost - Total*               $2,260      $2,160
 2.  Equipment Installation (25% of Item #1)            565         540
 3.  Piping Cost (20% of Item #1)                       452         432
 4.  Instrumentation (15% of Item #1)                    339         324
 5.  Electrical (10% of Item #1)                        226         216
 6.  Buildings (20% of Item #1)                         452         432
 7.  Site Preparation (10% of Item #1)                  226         216
 8.  Plant Design and Engineering (25% of Item #1)       565         540
 9.  Auxiliaries (30% of Item #1)   .                    678         648
                    PLANT COST                       $5,673      $5,508
10.   Contingency (15% of Plant Cost)                     864         826
                    TOTAL PLANT COST                 $6,627       $6,334
11.  Plant Start-up Costs                               125          125
12.  Interest During Construction  (10%)                  497          475
                    TOTAL  FIXED CAPITAL  COST          $7,249       $6,934
13.   Working Capital (10%  of fixed  total  cost)           725          693
                    TOTAL  CAPITAL  COST                $7,974       $7,627

 *Based on  data from Mehta and  0'

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                                   -86-
        TABLE XIX.  Operating Cost for Neutral Roasting Operation*
                                                        Cost $1000's/yr
                                                      MorenciPima
  1.   Operating Labor (6 men/shift)                    $  250      $  250
  2.   Maintenance Labor (3% of fixed capital costs)       217         208
  3.   Supervision (18% of Items 1 & 2)                     84          82
  4.   General Plant Overhead (2/3 of Items 1-3)           367         360
  5.   Power (6 kwh/ton feed)                               30          30
  6.   Natural Gas                                       Ij46o       1,400
  7.   Boiler Feed Water                                    25          26
  8.   Maintenance Supplies and Parts                      109         104
 9.   Operating Supplies                                   55          54
10.  Technical  Services (lab,  etc.)                       35          35
11.  Pyrite ($2/ton)                                      ^0         246
12.  Taxes and Insurance  (2%  of  plant cost)               132         127
                    TOTAL                             $2,854      $2,922
13.   Contingency (10% of Items  1-12)                     285         296
14.   Depreciation  (15 years)                             532         508
                    TOTAL  OPERATING  COST              $3,671      $3,726
     Cost  Per Pound  Copper Produced                   $0.0185     $0.0188
 *Based  on  data  from  Mehta and

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                                    -87-
      Capital  costs for Morenci  and Pima concentrates  were estimated to be
 $7,974,000 and $7,627,000,  respectively.   Yearly operating costs  were
 $3,671,000, and $3,726,000,  which  is  equivalent  to  $0.0185 and  $0.0188 per
 pound of copper produced.

 LEACHING OPERATION
      In  the leaching  circuit the neutral  roasted concentrates are reacted
 with  hydrochloric  acid  in a  series  of closed  leach  tanks  to produce H?S
 and dissolve  the iron.  The  leaching  temperature is 80  to 85C.   The solids
 residence time in  the  leaching  tanks  is assumed  to  be three hours.   Reaction
 temperatures  are maintained  using  low pressure steam  from the waste heat
 boilers.   Mechanical agitation  is  provided in each  leach  tank.  Liquid hold
 up in  the leach tanks  is about  200,000 gallons for  the  Morenci  system  and
 130,000  gallons for the Pima  system.  The hydrogen  sulfide  produced would
 be scrubbed with dilute hydrochloric  acid to remove any traces  of hydrogen
 chloride,  cooled to about 50C  and  sent to the SO-  recovery  system  (a portion
 of the H2S would be consumed  for copper recovery).  A nitrogen  or neutral
 flue gas  sweep would be used  on each  leach tank  to facilitate H?S handling
 in the off-gas system.  Nitrogen would be obtained from the oxygen  plant
 (required  for converter operation)  and would amount to about one-half the
 HpS on a  volume basis.
     The  leach slurry would be filtered through a rotary drum filter.  The
washed solids would pass to the converter.  The  leach solution  (plus wash)
would flow to a copper recovery tank where it would be sparged with H?S
 to precipitate any dissolved copper and precious  metals.  Liquid residence
 in the tank would be about 30 minutes.  The resulting slurry would be
filtered through a  primary rotary drum filter and a polishing filter.  The
washed solids would be sent to the  converter and  the clear leach solution
 (plus  wash) to the  acid recovery system.
     Estimating the capital  and operating costs  for the  leaching operation
is more subject to  question  than other sections of the process  because  the
leaching operation  has not been demonstrated  on a pilot  or plant scale.

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         TABLE XX.  Capital Cost for the Leaching Operation

                                                      Cost $1000's
                                                    MorenciPimT
  1.   Purchased Equipment Cost - Total               $ 2,700    $2,000
  2.   Equipment Installation (25% of Item #1)            675       500
  3.   Piping Cost (25% of Item #1)                       675       500
  4.   Instrumentation (15% of Item #1)                   405       300
  5.   Electrical (10% of Item #1                         270       200
  6.   Buildings (20% of Item #1)                         540       400
  7.   Site Preparation (10% of Item #1)                  270       200
  8.   Plant Design and Engineering (25% of Item #1)       675       500
  9.  Auxiliaries (30% of Item #1)                       810       600
                 PLANT COST                         $ 7,020    $5,200
10.  Contingency (20% of Plant Cost)                   1,404     1,040
                 TOTAL PLANT COST                   $  8,424     $6,240
11.  Plant Startup Cost                                 150        120
12.  Interest During Construction (10%)                  632        468
                 TOTAL FIXED CAPITAL  COST            $  9,206     $6,828
13.   Working Capital  (10% of fixed capital  cost)         921       .683
                 TOTAL  CAPITAL  COST                  $10,127    $7,511

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                                -89-
       TABLE XXI.   Operating Costs  for  the  Leaching Operation

                                                   Cost  $1000's/yr
                                                  Morenci      Piiria
  1.   Operating Labor  (5  men/shift)                   208       208
  2.   Maintenance  Labor                               276       205
  3.   Supervision  (18% of items  1 and 2)               87        74
  4.   General  Plant Overhead  (2/3 of Items  1-3)       381        325
  5.   Power                                           200        140
  6.   Process  Water                                    30         30
  7.   Cooling  Water                                    40         49
  8.   Steam  (available from waste heat boilers)              	
  9.   Hydrochloric Acid (0.1% of inventory lost/day)    40         27
10.   Maintenance Supplies and Parts                  138        102
11.  Operating Supplies                                57         49
12.  Technical Services                               100        100
13.  Taxes and Insurance                              168        125
                 TOTAL                            $1,725      $1,425
14.   Contingency (15% of Items  1-12)                  259         214
15.   Depreciation (15 years)                          675         501
                 TOTAL  YEARLY  OPERATING  COST      $2,659     $2,140
                 COST PER  POUND  OF  COPPER         $0.0134    $0 0108
                 PRODUCED

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                                     -90-
 Equipment costs  and labor requirements  were  estimated  from  data  for  various
 mineral  leaching operations  reported  in the  literature,  allowing for inflation
 and a chloride-containing system.   Capital and yearly  operating  costs  for  the
 Morenci  and Pima systems  are summarized in Tables  XX and XXI.  The capital
 costs are estimated to  be $10,130,000 and $7,510,000 for the  two cases.
 The lower capital  cost  for the  Pima system results  from  the higher acid
 concentration  used and  the reduced  volume of liquid which is  handled.  The
 yearly operating costs  for the  two  cases are estimated to be  $2,660,000
 and $2,140,000,  which corresponds to  a  cost  $0.0134/lb of copper for the
 Morenci  system and $0.0108/lb of copper for  the Pima system.

 ACID RECOVERY  SYSTEM
      In  the  acid recovery  system the  leach solution (plus wash)  is treated
 to  regenerate  the  hydrochloric acid and convert the ferrous chloride to
 ferric oxide.  The  system  proposed  for  use is the Lurgi  System,'4' which
 is  used  for  the  regeneration of hydrocloric  acid pickle  liquor.   In  the
 Lurgi-type system  the leach  solution  is first concentrated  in an evaporator.
 The  concentrated  liquor is then fed to a fluid bed reactor where the HC1 and
 water are vaporized and the  ferrous chloride converted to ferric oxide.

                   2FeCl2 + 2H20 + 1/202 	> 4HC1  + Fe^

 The  reactor  is operated at 800C and  is gas  fired.   The hot off-gas from the
 reactor passes through a cyclone to remove Fe^O- particulates and then to
 the evaporator for heat recovery.   The cooled gases from the evaporator
 pass to an absorber where the HC1  is absorbed to form 4 to 5t hydrochloric
 acid, which  is recycled to the leach circuit.  The  HC1  free gases are then
 vented to the stack.  The ferric oxide is obtained  as  chloride-free,  free-
 flowing,  coarse,  granular solid having a bulk density  of about 150 lb/ft3.
     Capital and  operating costs for the acid recovery  system are summarized
 in Tables XXII and XXIII.   Operating costs  were  calculated using  the  data
from Table XVI.  Capital costs for  the Morenci and  Pima systems are estimated

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                                    -91-








            TABLE XXII.  Capital Costs for Acid Recovery System





                                                             Cost $1000's
                                                          Morenci     n'ma
 1.  Purchased Equipment Cost - Total*                    $ 4,100    $ 3,100



 2.  Equipment Installation (25% of Item#l)                1,025        775



 3.  Piping Cost (25% of Item #1)                           1,025        775



 4.  Instrumentation (15% of Item #1                           615        465



 5.  Electrical (10% of Item #1)                              410        310



 6.  Buildings (20% of Item #1)                               820        620



 7.  Site Preparation (10% of Item #1)                        410        310



 8.  Plant Design and Engineering (25% of Item #1)           1,025        775



 9.  Auxiliaries (30% of Item #1)                           1,230   .     930
                 PLANT COST                               $10,660    $ 8,060



10.  Contingency (15% of plant cost)                         1,600      1,210
                 TOTAL PLANT COST                         $12,260    $  9,270



11.  Plant Startup Costs                                      200        175



12.  Interest During Construction (10%)                        920        695
                 TOTAL FIXED CAPITAL COST                 $13,380    $10,140



13.   Working Capital  (10% of fixed capital  cost)               134        101
                 TOTAL CAPITAL  COST                       $13,514    $10,241

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                                    -92-
           TABLE XXIII.  Operating Costs for Acid Recovery System

                                                            Cost $1000's/yr
                                                           MorenciPima
 1.  Operating Labor                                          210        210
 2.  Maintenance Labor (3% of fixed capital  costs)            401        304
 3.  Supervision (18% of Items 1  and 2)                       110         93
 4.  General  Plant Overhead (2/3  of Items 1-3)                 483        405
 5.  Power                                                     50         35
 6.  Natural  Gas                                            4,950      3,300
 7.  Cooling Water                                             60         40
 8.  Process  Water                                            100         65
 9.  Hydrochloric Acid (22 Be1)                                80         60
10.  Maintenance Supplies and Parts                           200        152
11.  Operating Supplies                                        72         61
12.  Technical Services                                        40         40
13.  Taxes and Insurance                                       245        185
                 TOTAL                                      $7,001      $4,950
14.   Contingency (10% of Items  1-13)                           700         495
15.   Depreciation                                             901         683
                 TOTAL  OPERATING  COST                       $8,602      $6,128
                 COST PER  POUND COPPER  PRODUCED             $0.0434     $0.0309

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                                    -93-
 to be  $13,500,000  and  $10,200,000,  respectively.  The yearly operating costs
 are estimated  at $8,600,000 and $6,100,000, corresponding  to $0.043 and
 $0.031  per  pound of  blister copper  produced.
     Acid recovery systems of the size required to treat the plant leach
 liquor have not been built to date.  There is, therefore,  some question as
 to scale-up from existing units.  In addition, there is only limited operat-
 ing data available on  the Lurgi system.  For these reasons, conservative
 estimates were used  for both capital costs and operating costs.  By a more
 detailed study of  the  Lurgi and other systems such as that of Haveg,'4' it
 should  be possible to  reduce both capital and operating costs significantly.
     Natural gas for heating the reactors and evaporators accounts for more
 than 60% of the yearly operating costs.  By improved design, additional heat
 recovery (multiple effect evaporators) etc., it should be possible to reduce
 fuel requirements  by a significant amount.

 S00 RECOVERY SYSTEM
 		
     The citrate process developed by the U.S.  Bureau of Mines for recovery
 of  S09  involves the reaction of S00 and H0S in an aqueous solution to form
                 (34)                    ^
 elemental sulfur.     '
                         S02 + 2H2S -> 3S + 2H20
 The basic steps in the process are:
 1.  Washing of the cooled S02-bearing gas to remove particulates and SO.,,
 2.  absorption of the S02 from the cleaned gas by a solution of citric
     acid,  sodium  citrate and sodium thiosulfate,
 3.  reaction of the absorbed S0? with H?S in a closed vessel,  precipitating
     elemental  sulfur and regenerating the  absorption solution, and
 4.  recovery of the sulfur by oil  melting.
 S02 recoveries  of 90 to 99% have been reported.^   '   In this study a
 recovery of 98% was assumed.   A small amount of the  S02 is  converted  to
 sulfate in the  process.  For this study it  was  assumed that enough fkS
would be required  to react with all  of the  SOp  absorbed,  ignoring any
 possible sulfate formation.

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                                     -94-
      The citrate process has only been demonstrated on a small  pilot-plant
 scale.   Scale up to the size required for this study may be questionable
 until  the plant scale operation,  which is underway at the Bunker Hill
 Smelter, Kellogg, Idaho, has been demonstrated.   If a lower S0? recovery
 could  be tolerated (90 to 95%)  then  a modified Claus process might be  used.
 The Claus process has been demonstrated .in large-scale applications  and
 would  have lower capital  and operating costs  than  the citrate process.
     Capital  and operating costs  for S02  recovery  using the citrate  process
 were estimated  using  Bureau of  Mines data reported  in 1973^36^  with  an
 allowance for inflation.   It was  assumed  that the  gas flow  to the  S0?
 recovery unit would be approximately 30,000 SCFM with an average S0? con-
 centration of about 6%.   To account  for fluctuations  in gas flow the system
 was  sized to  handle 38,000 CFM.   Capital  and  yearly  operation costs  for
 each case under  consideration are summarized  in Tables  XXIV and XXV.  Esti-
 mated capital costs for  the Morenci  and Pima  systems  are  $8,270,000  and
 $8,620,000, respectively,  while yearly  operating costs  are  estimated to be
 $2,100,000 and $2,180,000.   The corresponding  costs per pound of copper
 produced  are  $0.0106  and  $0.0110.

 CONVERTER  OPERATION
     The  copper-bearing solids from  the leach  circuit are fed directly  to
 a converter.  This  feed material  is  high  in copper (^50 wt%) and low in
 iron (510 wt% Fe).  The sulfur content  (^20 wt%) is also lower than in  the
 normal  converter feed.  For  this  reason, much of the heat required for
 converted operation must be  supplied by natural gas or other fuel.   A portion
 of the oxygen required for reaction with the sulfur and burning  of the
 natural  gas will be supplied from an oxygen plant (^20%).  The rest will be
 supplied by air.  The off-gas from the converter will pass through a waste
 heat boiler and then through cleaning equipment to remove the bulk of the
 particulates and cool  the gas to about 300C.   The gases then pass  to the
S0? recovery unit.

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                                    -95-
              TABLE XXIV.   Capital  Costs  for  S02  Recovery System*

                                                             Cost $1000's
 *Based on  data  from J.  B.  Rosenbaum,  et.
                                                           Morenci    Pima
  1.   Purchased Equipment Cost - Total                      $2,300    $2,400
  2.   Equipment Installation (25% of Item #1)                  575       600
  3.   Piping Cost  (25% of Item #1)                             575       600
  4.   Instrumentation (15% of Item #1)                         345       350
  5.   Electrical (10% of Item #1)                              230       240
  6.   Buildings (20% of Item #1)                               460  '     480
  7.   Site Preparation (10% of Item #1)                        230       240
  8.   Plant Design and Engineering (25% of Item #1)            575       600
  9.   Auxiliaries (30% of Item #1)                             690       720
                 PLANT COST                                $5,980    $6,240
10.  Contingency (15% of plant cost)                          897       936
                 TOTAL PLANT COST                          $6,877     $7,176
11.  Plant Startup Costs                                      125        125
12.  Interest During Construction (10%)                        516        538
                 TOTAL FIXED CAPITAL  COST                   $7,518     $7,839
13.   Working Capital  (10% of fixed  capital  cost)               752        784
                 TOTAL  CAPITAL  COST                         $8,270     $8,623

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                                     -96-
             TABLE  XXV.  Operating Costs for S02 Recovery System
                                                            Cost  $1000's/yr
                                                           MorenciPima
 1.  Operating Labor (4 men/shift)                          $   167      $   167
 2.  Maintenance Labor                                        226         235
 3.  Supervision (18% of Items  1  and 2)                         71          73
 4.  General  Plant Overhead (2/3  of Items  1-3)                 309         317
 5.  Power                                                     20          21
 6.  Natural  Gas                                               50          52
 7.  Steam (waste steam available at no  charge)                        
 8.  Process  Water                                             15          16
 9.  Cooling  Water                                             35          36
10.  Chemicals                                                162         169
11.  Maintenance Supplies  and Parts                           113         118
12.  Operating Supplies                                        46          46
13.  Technical Services                                        60          60
14.  Taxes and Insurance                                      138         144
                 TOTAL                                     $1,412     $1,454
15.   Contingency (10% of Items  1-14)                           141        145
16.   Depreciation (15 year)                                    551        575
                 TOTAL  OPERATING  COST                      $2,104     $2,176
                 COST  PER  POUND COPPER  PRODUCED            $0.0106    $0.0110

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                                     -97-
      The  copper  leaves  the converter at a purity of about 98.8% and contains
 about 0.2%  sulfur.  The blister copper is transferred to a holding furnace
 and  then  to a  refining  furnace.  The copper from the refining furnace is
 cast into anodes  for further refining.  The off-gases from the holding and
 refining  furnaces are vented to the stack.  Approximately 0.1% of the
 sulfur in the  original  concentrates would be present in the off-gas as S0?.
 The  cost  estimates include all operations through the casting of anodes but
 does  not  include  subsequent operations.
      The  slag  from the  converter contains a small but significant amount of
 copper.   This  slag would be recycled into the concentrator-flotation plant
 where  the copper would  be recovered as part of the copper concentrate.
     The  steam produced in the waste heat boiler meets, most of the steam
 requirements of the entire plant operation.   Additional  steam capacity
 required  is minimal.
     The  capital and operating costs for the converter operation are sum-
 marized in Tables XXVI  and XXVII.   Equipment and operating cost data were
 obtained  from a 1973 Bureau of Mines publication by Bennett, et al.'38'
 Capital costs for the Morenci and  Pima systems are $12,200,000 and
 $12,700,000, respectively.  Yearly operating costs are estimated to be
 $6,120,000 and $6,260,000, which is equivalent to $0.0309 and $0.0316 per
 pound of  anode copper.

TOTAL PROCESSING COSTS
     The  total estimated capital  and yearly  operating costs  for a 300 ton/day
copper smelter are summarized in Tables XXVIII and XXIX.   The capital  costs
for the Morenci and Pima cases are $52,000,000 and $46,700,000,  respectively.
These estimates include all  of the facilities  required to convert copper
concentrate into copper anodes.
     The yearly operating  costs  for the two  cases are estimated  at
$23,150,000 and $20,430,000,  which corresponds to a  cost  of  $0.117 per
pound of anode copper for  the Morenci  case and $0.103 for the Pima case.

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                                      -98-

              TABLE  XXVI.   Capital Costs  for Converter Operation

                                                             Cost $1000's
                                                           Morenci     Pima
 1.  Purchased Equipment Cost - Total*                     $.3,400    $ 3,500
 2.  Equipment Installation (25% of Item #1)                   850        875
 3.  Piping Cost (20% of Item #1)                              680        700
 4.  Instrumentation (15% of Item #1)                           510        525
 5.  Electrical (10% of Item #1)                               340        350
 6.  Buildings (25% of Item #1)                                 850        875
 7.  Site Preparation (10% of Item #1)                         340        350
 8.  Plant Design and Engineering (25% of Item #1)              850        875
 9.  Auxiliaries (30% of Item #1)                            1,020      1,050
                 PLANT COST                                $  8,840     $  9,100
10.  Contingency (15% of plant cost)                          1,326       1,365
                 TOTAL PLANT COST                          $10,166    $10,465
11.  Plant Startup Costs                                        175         175
12.  Interest During Construction (10%)                         762         785
                 TOTAL  FIXED CAPITAL  COST                  $11,103    $11,425
13.   Working Capital  (10% of fixed  capital  cost)              1,110       1,143
                 TOTAL  CAPITAL  COST                        $12,213    $12,668

 *Based on  data  from Bennett, et.al.^38^

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                                     -99-
            TABLE  XXVII.  Operating Costs for Converter Operation*
                                                            Cost $1000's/yr
                                                           MorenciPiiria
 1.  Operating Labor (28 men/shift)                         $1,165    $1,165
 2.  Maintenance Labor                                        340        345
 3.  Supervision (18% of Items 1  and  2)                        271         272
 4.  General Plant Overhead (2/3  of  Items  1-3)               1,184      1,188
 5.  Power                                                    160        162
 6.  Natural Gas                                              750        755
 7.  Cooling Water                                             30         30
 8.  Boiler Water                                              30         30
 9.  Flux-Converter ($5.50/ton)                                 91         163
10.  Refractory                                               150        150
11.  Maintenance Supplies and  Parts                            167        171
12.  Operating Supplies                                        178        178
13.  Technical Services                                        100         100
14.  Taxes and Insurance                                       203     .    209
                 TOTAL                                     $4,819     $4,918
15.   Contingency (10% of Items  1-141                          482        492
16.   Depreciation (15 years)                                  814        845
                 TOTAL  YEARLY OPERATING COSTS              $6,115     $6,255
                 COST PER  POUND OF  COPPER ANODE            $0.0309    $0.0316
 *Based on  data  from Bennett, et. al.
(38)

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                                      -100-
         TABLE XXVIII.   Capital  Costs  for  a  300 Ton/Day Copper Smelter


                                                             Cost  $1000's
                                                           MorenciPi ma
 1.  Neutral  Roasting Operation                             $  7,974    $ 7 627

 2.  Leaching Operation                                     10,127      7 511

 3.  Acid Recovery  System                                  13,514     10 241

 4.  S02  Recovery Unit                                        8>270      8>623

 5.  Converter Operation                                    12 213     12 668
                TOTAL                                     $52,098    $46,670
    TABLE XXIX.  Yearly Operating Costs for a 300 Ton/Day Copper Smelter
                                                           Cost $1000's/yr
                                                          Morenci     Pima
 1.  Neutral  Roastino  Operation                            $ 3^71    $ 3 726

 2.  Leaching Operation                                      2,659      2 140

 3.  Acid Recovery System                                    8,602      6 128

 4.  S02 Recovery Unit                                       2,104      2 176

 5.  Converter Operation                                     6,115      6 255
                TOTAL OPERATING COSTS                     $23,151     $20,425

                COST PER POUND OF ANODE COPPER            $0.117     $0.103

6.  Credit for Sulfur ($20/ton)                             3546i       3^35

7.  Credit for Fe203 ($10/ton)                              1,757       -\ j776
                COST PER POUND OF  ANODE  COPPER             $0.091      $0  078
                (with credit  for S and Fe0

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                                    -101-
 If  credit  is allowed for the sulfur  ($20/ton) and ferric oxide ($10/ton),
 the operating costs are reduced to $0.091 and $0.078 for the Morenci and
 Pima  cases, respectively.
      Although for comparison it would be most convenient to have an operat-
 ing,  elemental sulfur-producing, state-of-the-art S0? air pollution abate-
 ment  system integrated with a conventional copper smelter, such a combina-
 tion  does  not, to our knowledge, exist at this time.  The extensive study
 made  by the Fluor Utah Corporation and as authoritatively reviewed by
 Swan*   '   ' is therefore taken as the basis for reference for this compari-
 son.  From that study which covers the entire U.S. copper smelting industry,
 the installation cost for 90% emissions control for an add-on system result-
 ing in elemental sulfur production would amount to $415,000,000 (1970 costs).
 The annual operating cost for these additional facilities without credit for
 by-product sulfur was reported to be $130,000,000, which would amount to an
 additional cost for the production of copper of 4-1/3 cents per pound, based
 on  this 1970 study estimate.  A reasonable addition to this cost of 1 to
 1.5 cents per pound should be added to account for inflationary factors
 since the 1970 estimate was made.   It is likely then that at present and
 near-future costs the addition should be in the range of 5 to 6 cents per
 pound of copper for emissions control as provided for in the Fluor study.
A similar situation should be projected for overall  smelting costs.   A 1973
                           f 38)
U.S. Bureau of Mines reportv  '  estimated the overall  smelting cost to be
6.4 cents per pound of copper.   With anticipated near-future increases in
 fuel and labor costs, this figure  should probably also be adjusted upward to
a range of 7 to 8 cents.   Thus,  by combining these cost figures,  a figure
of 12 to 14 cents is obtained for  present and near-future costs for copper
smelting with equivalent  sulfur dioxide air pollution abatement including
recovery of some portion  of the  sulfur in elemental  form.   This range can
be compared with the 10 to 12 cents range for the proposed process concept
with a tendency to favor  the lower figure.   These results indicate that the
proposed process concept  has a  favorable economic outlook when viewed under
circumstances  which appear to be conservative.

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                                    -102-
                                CONCLUSIONS

ADEQUATE  PRODUCTION OF HYDROGEN SULFIDE FOR ELEMENTAL SULFUR RECOVERY
IS ACHIEVABLE UITHOUT NEED FOR ANY SPECIFIC CHEMICAL REDUCTANT
     The  proposed process concept can result in the production of sufficient
hydrogen  sulfide from pyritic copper concentrates for reaction with sulfur
dioxide emissions from converters for elemental sulfur recovery without
need for  any specific chemical reductant.  The process may be able to use
any type  of fuel.

THE ECONOMICS OF AN OVERALL SMELTING PROCESS UTILIZING THE PROCESS
APPEARS FAVORABLE
     The  preliminary economic assessment indicates that an overall smelting
process employing the concept has favorable costs for achieving copper
production with adequate sulfur emissions control.  Credit for the iron
oxide and elemental  sulfur provide an even more favorable economic picture
at conservative values.   These preliminary evaluations suggest an adequate
basis for continued support on the development of the process through the
pilot plant design phase by EPA and industry.

NEUTRAL ROASTING IS A FEASIBLE MEANS OF CONVERTING THE IRON SULFIDE TO THE
ACID-SOLUBLE FORM
     The results of this study show that neutral  roasting is a chemically
feasible means of assuring the maximum selective  solubilization of the iron
sulfide in pyritic copper concentrates without major dissolution of copper.
Neutral-roasting as  studied by others for pyrite  alone supports the proposed
process as a feasible and practical  method for also treating the pyritic
copper concentrates  for  preliminary removal  of a  significant fraction of the
sulfur initially as  elemental  sulfur as  well  as for converting the remain-
ing iron sulfide into a  soluble  form.

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                                    -103-
 HYDROCHLORIC  ACID  LEACHING  IS  A  PRACTICAL METHOD  FOR  HYDROGEN  SULFIDE
 PRODUCTION
      Hydrochloric  acid  leaching  of  neutral-roasted concentrates  is a direct
 and  practical  means  for selective dissolution of  iron and  release of
 hydrogen  sulfide.  Copper,  which may be dissolved in minor amounts, is
 recoverable to a major  extent  along with any solubilized precious metals
 (silver)  by simple treatment of  the cooled solution with the hydrogen sulfide,
 Hydrogen  sulfide can be produced by this means  in amounts  sufficient for
 reaction  with  all  of the sulfur  dioxide released  in the step of  converting
 the  enriched  copper  sulfide to blister copper.  Known present  state-of-the-
 art  technology is  applicable to  conduct the hydrogen sulfide-sulfur dioxide
 reaction  for  high-purity elemental sulfur production via the so-called Claus
 process or by  the  U.S.  Bureau  of Mines Citrate  process.

 HYDROCHLORIC ACID  REGENERATION IS A PRESENT STATE-OF-THE-ART PROCESS
     Hydrochloric  acid  regeneration from acid-ferrous chloride solutions
 is regarded as a present  state-of-the-art process which does not need demon-
 stration at the laboratory  scale.  Existing processes which are already
 employed on a  large  scale in the steel  pickling industry can accommodate
 ferrous chloride solutions of any concentration and acidity.

 THE BLISTER COPPER WILL  BE RECOVERED IN SATISFACTORY YIELD AND QUALITY
     The process is  essentially a closed one insofar as copper processing is
 concerned.  Solids are processed by present state-of-the-art methods  in
 conventional  converters,  liquids are specially treated for very effective
 copper recovery.  Dusts would be retained by the usual methods.  If anything,
 a higher yield of copper would be anticipated because of the avoidance of
 the usual  reverberatory slag losses.  Quality should be essentially no dif-
 ferent from the usual blister copper.   The small amounts of converter slag
would be recycled to the concentrator for additional  cop'per recovery.   In
conventional  smelting this slag is  added  instead to  the reverberatory
furnaces.

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                                    -104-
 THE IRON  OXIDE  AND  HYDROGEN  SULFIDE DERIVED SULFUR WILL BE OF HIGH QUALITY
      The  iron oxide by-product  produced during hydrochloric acid regeneration
 from ferrous chloride  solutions  is expected to be of high quality and suit-
 able for  marketing  for steel-making.  Small amounts of other metal chlorides
 which may also  be oxidized,  such as zinc, principally, and possibly lead,
 are expected to be  of  no consequence in determining the suitability of the
 iron oxide.  The sulfur obtained via the hydrogen sulfide-sulfur dioxide
 is  expected to  be of the very high purity as represented by that typically
 produced  via the Glaus process.  The sulfur produced in the neutral roasting
 operation  will  be relatively impure but can be purified to suitable quality
 by  present state-of-the-art processes/26^

 SIGNIFICANT ECONOMIES APPEAR TO BE LIKELY FOR THE PROCESS
      Although it is felt that the preliminary economic evaluation of the
 process concept is adequate to encourage continued consideration and support
 for  development of an overall process,  significant economies  appear to be
 probable by more intensive consideration of alternative lower-cost fuels and
 engineering refinements to assure increased heat economy.   The estimates,
 for  convenience and  simplicity  at this stage of the  investigation,  assumed
 the  use of high-cost natural  gas as the fuel.

THE  PROPOSED PROCESS IS APPLICABLE  IN ANY  SULFIDE SMELTER  HAVING SUFFICIENT
PYRITE
     Given a sufficient supply of pyrite,  any  operator  of  a conventional
smelter for copper sulfides concentrates could consider the installation  of
the proposed process.   Such installation may  be more readily  incorporated
into a new smelter than in  an old one.   Space  and physical arrangement would
be the principal problems  in  modifying  old  smelters  to  accommodate  the
revised process.

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                                    -105-
 INCREASED  COPPER AND  PRECIOUS METAL RECOVERY FROM THE ORE BODY SHOULD BE
 POSSIBLE
      By use  of  large  amounts of pyrite which is normally discarded, the
 small  but  significant amounts of copper and accompanying precious metals
 in  such discard would now be almost completely recovered at essentially
 very  low incremental  cost.  For exploitation of some ore bodies the
 increased  recoveries  may be quite significant.

 OTHER  BENEFITS  IN RECOVERY OF METAL VALUES, RESOURCE CONSERVATION AND
 ECONOMICS MAY BE REALIZED
     Because of the need for pyrite, a lower-grade ore may be beneficially
 utilized.   The use of the major slag-forming minerals such as silica and
 limestone would be greatly reduced.   The handling and disposal of the large
 amounts of useless slag would be greatly reduced.   The two other major
 components of the concentrates,  iron and sulfur,  would be recovered in
 suitable purity such that beneficial  utilization  could then  be anticipated
with significant economic credit to  the overall  process.   The large-volume,
 low-strength sulfur dioxide  flue gases characteristic  of reverberatory
furnace operations  would be  avoided.

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                                    -106-
                             RECOMMENDATIONS

     In view of the demonstrated chemical  feasibility,  favorable economics
and indicated other significant benefits which  appear to  be probable  by  the
operation of the proposed process concept,  the  following  recommendations in
support of continued development of the  integrated  process  appear appropriate:
 1.  Large bench-scale experimental  work should be  continued to  obtain
     needed data on dynamic  systems  suitable  for the  design of a pilot plant.
     Such work would be done on the  steps  of  neutral  roasting, hydrochloric
     acid leaching, residual  copper  precipitation and separation of solid,
     enriched copper sulfide and residues  from  the  ferrous  chloride leach
     liquors.
 2.  Work jointly with an appropriate  architect-engineer-constructor  to
     assure that information  suitable  for  the design  and  estimating of the
     acid regeneration system is obtained  in  the continued  laboratory studies.
 3.  Conduct engineering  studies to  determine the most  economical  fuel to be
     used and equipment refinements  which may be justified  to assure most
     beneficial  heat economy.
 4.  Refine the  economic  assessment  of the overall  process.
 5.  Conduct process  engineering and economic comparisons between  this
     process and  other copper smelting processes which  exist in  the U.S. or
     are  being considered  for the production of elemental sulfur.

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                                    -107-
                                REFERENCES


  1.   Levy,  S.  I.  and G. W. Gray, British Patent 309,269 (1928), "Treating
      Pyrite Residues," and French Patent 668,093 (1928), "Recovering
      Copper."

  2.   Levy,  S.  I., U. S. Patent 1,746,313 (1930), "Treatment of Copper-Rich
      Material," and also U.S. Patent 1,980,809 (1934), "Production of
      Ferric Oxide and Other Metal Values from Pyrites."

  3.   Parsons,  H. W. and T. R. Ingraham, "The Hydrogen Sulfide Route to
      Sulfur Recovery from Base Metal Sulfides, Part I, The Generation of
      H^S  from  Base Metal Sulfides," June 1970.

  4.   Reeve,  D. A. and T. R. Ingraham, "The Hydrogen Sulfide Route to Sulfur
      Recovery  from Base Metal Sulfides, Part III, Recovery of Iron Products
      from Ferrous Chloride Solution," Canada, Department of Energy, Mines
      and  Resources, Mines Branch, Ottawa, Mines Branch Program on Environ-
      mental  Improvement.

  5.   Cabri,  L. J., "New Data on Phase Relations in the Cu-Fe-S System,"
      Econ.  Geol., 68, pp. 443-54, 1973.

  6.   Cabri,  L. J., et al., Lau.  Mineral, ]2_, Pt 1,  1973.

  7.   Cabri,  L. J.  and S. R. Hall, "Mooihoekite and  Haycockite, Two New
      Copper-Iron Sulfides and Their Relationship to Chalcopyrite and
      Talnakhite," Am.  Mineral, Etf,  pp.  689-708, 1972.

  8.   Cabri, L. J.  and D.  C. Harris,  "New Compositional  Data on Talnakhite,"
      Econ. Geol.,  66,  pp.  673-75,  1971.

  9.  Clark, A. H., "An Unusual Copper-Iron  Sulfide,"  ibid,  65, pp.  590-91,
      1970.                                                 

10.  Hall, S. R.  and E.  J.  Gabe,  "The Crystal  Structure of  Talnakhite,"
     Am. Mineral,  57_,  pp.  368-80,  1972.

11.  MacLean, W.  H.,  et  al.,  "Exsolution Products  in  Heated Chalcopyrite,"
     Can.  I  Ea. Sci.,  9_,  pp.  1305-17, 1972.

12.  Schlegel, H.  and  A.  Schuller, Met  Kunde.  4, pp.  421-29,  1952.

13.  Merwin, H. E.  and  R.  H.  Lombard, "The  System Cu-Fe-S," Econ.  Geol.,
     32, pp. 203-84,  1937.                                   	

14.  Donnay, G and G.  Kullerud,  "High Temperature Chalcopyrite," Carnegie
     Inst. Washington  Year Book,  57^  p.  246,  1958.

15.  Roseboom, E.  H.,  Jr.  and  G.  Kullerud,  "The Solidus in  the System
     Cu-Fe-S Between  400  and  800C,"  ibid,  pp.  222-27, 1958.

16.  Brett,  P.  R.,  "The  Cu-Fe-S System," Carnegie Inst. Washington  Year,
     62, pp. 193-96,  1963.

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                                    -108-
 17.   Brett,  P.  R.,  "Experimental  Data  from the System Cu-Fe-S and Their
      Bearing on Exsolution  Textures  in Ores," Econ. Geol., 59, pp. 1241-69,

 18.   Yund, R. A.  and  G.  Kullerud,  "Thermal Stability of Assemblages in the
      Cu-Fe-S System,"  J.  Petrology,  7_, pp. 454-488, 1966.

 19.   Von  Gehlen,  K. and  G.  Kullerud, "Pyrrhotite-Pyrite-Chalcopyrite
      Relations,"  Carnegie Inst. Washington Year Rnnk. fil. nn  IZA.^

20.
21.
22.
23.

1962.
Kullerud, G.
Kullerud, G.
Kullerud, G.
Morimoto, N.
Acta Cryst. ,

, "The CugS5-Cu5FeS4 Join," ibid, 59_, pp. 114-16, 1960.
, "The Cu-S System," ibid, 59, pp. 110-11, 1960.
, "The Cu-Fe-S System," ibid, 63, pp. 200-202, 1964.
, "Structures of Two Polymorphic Forms of Cu^eS.,"
17, pp. 351-60, 1964. 5 4
 24.   Yund, R. A. and G. Kullerud, "The Cu-Fe-S System," Carnegie Inst.
      Washington Year Book, 59, pp. 111-14, 1960.

 25.   Yund, R. A. and G. Kullerud, "The System Cu-Fe-S," ibid, 60, pp.  180-
      81,  1961.

 26-   	 "Outokumpu Process for the Production of Elemental Sulfur
      from Pyrite," Sulfur, No. 50, 1964.

 27.   Kunda, W., V. N. Mackiw and B.  Rudyk, "Iron and Sulfur from Sulfidic
      Iron Ores," The Canadian Mining and  Metallurgical  (CIM)  Bulletin  for
      July, 1968, pp. 819-835.	

 28.   Mehta, B. R.  and P. T.  O'Kane,  "Economics of Iron  and Sulfur Recovery
      from Pyrites," ibid,  pp. 836-845.

 29.   Watkinson, A. P. and  C.  Germain, "Thermal Decomposition  of Pyrite in
      Fluid Beds,"  Noranda  Research Centre, Pointe Claire,  Quebec, Canada,
     A paper presented at  the 100th  AIME  Annual  Meeting, New  York City,  NY
     March 1-4, 1971.

 30.  Tkachenko, 0. B., A.  L.  Tseft,  and 0. Kunseitov,  "Thermal  Decomposition
     of Pyrite in  Vacuum and  Subsequent Treatment of the Residue,"  Tr.  Inst.
     Met. Obogashch., Akad.  Nauk Kaz., SSR,  19,  pp.  53-58, 1966.

 31.  Tkachenko,  0.  B., A.  L.  Tseft,  I. G.  Dem'yanikov,  and V.  Ya.  Slashchinina,
      "Thermal  Decomposition  of Chalcopyrite  in Vacuum and  Subsequent Treatment
     of the Residue," Tr.  Inst.  Met.  Obogashch.,  Akad.  Nauk Kaz.  SSR,  19,
     pp. 41-52,  1966.                                                 

32.  Isakova,  R. A., M. I.  Usanovich,  N. A. Potanina  and L.  E.  Ugryumova,
      (Inst.  Met. Obogashch, Alma-Ata, USSR),  Izv.  Akad. Nauk  Kaz..  SSR,
     Ser. Khim,  ], (5), pp.  78-81,  1969.

33.  Van Weert,  G., K.  Mah and N.  L.  Piret,  "Hydrochloric  Acid  Leaching of
     Nickeliferous  Pyrrohotites  from  the  Sudbury  District," CIM Bulletin,
     No.  1,  pp.  97-103,  1974.                               	

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                                    -109-
34.  Ingraham, T. R., H. W. Parsons and L.  J.  Cabri,  "Leaching of
     Pyrrhotite with Hydrochloric Acid," Can.  Metal.  Quart., 11,  pp  407-11,
     1972.                               	3	   

35.  Thornhill, P. G., E. Wigstal and G. Van Weert,  "The Falconbridge Matte
     Leach Process," J. Metals. 23, No. 7,  pp.  15-18, 1971.

36.  Rosenbaum, J. B., W. A.  McKinney, H.  R. Beard,  L.  Crocker and
     W. I. Nissen, "Sulfur Dioxide Emission Control  by Hydrogen Sulfide
     Reaction in Aqueous SolutionThe Citrate  System," Bureau of Mines,
     U.S. Department of the Interior, Report of Investigations; 1973,
     RI-7774.	

37.  Harvey, A. E., et al., "Simultaneous  Spectrophotometric Determination
     of Iron (II) and Total Iron with 1,10-Phenanthroline,"  Anal.  Chem.,
     27, pp. 26-29, 1955.                                   	

38.  Bennett, H. J., L. Moore, L. E.  Uelborn and  J.  E.  Toland, IC-8598,
     "An Economic Appraisal of the Supply  of Copper  from Primary  Domestic
     Sources," U.S. Bureau of Mines Information Circular/1973.

39.  PB-208-293, "The Impact  of Air Pollution Abatement on the Copper
     Industry," by Fluor Utah Engineers and Contractors,  Inc., for the
     Kennicott Copper Corporation, April 20, 1971.

40.  Swan, D., "Study of Costs for Complying with Standards  for Control of
     Sulfur Oxide Emissions from Smelters," Mining Congress  Journal,  vol. 57,
     No. 4, pp.  76-84,  April  1971.

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                          	-nn-	
                                 TECHNICAL REPORT DATA
                          (Please read luu/uctiont on the reverse before completing)
 1. REPORT NO.
  EPA-650/2-74-085-b
                            2.
                                                       3. RECIPIENT'S ACCESSION>NO.
 4. TITLE ANOSUBTITLE
  Control of Sulfur Dioxide Emissions from Copper
   Smelters; Volume II--Hydrogen Sulfide Production
   from  Copper Concentrates	
                                   5. REPORT DATE
                                    September 1974
                                   6. PERFORMING ORGANIZATION CODE
 7. AUTHOR(S)
                                                       8. PERFORMING ORGANIZATION REPORT NO
  C. A. Rohrmann and H. T. Fullam
 9. PERFORMING ORG MMIZATION NAME AND ADDRESS
  Battelle Pacific Northwest Laboratories
  Battelle Boulevard
  Richland, Washington  99352
                                                       10. PROGRAM ELEMENT NO.
                                    1AB013: ROAP 21ADC-056
                                   11. CONTRACT/GRANT NO.
                                    68-02-0025
 12. SPONSORING AGENCY NAME AND ADDRESS
  EPA, Office of Research and Development
  NERC-RTP, Control Systems Laboratory
  Research Triangle Park, NC 27711
                                   13. TYPE OF REPORT AAIO PERIOD COVERED
                                    Final; 6/73-4/74	
                                   14. SPONSORING AGENCY CODE
 15. SUPPLEMENTARY NOTES
 16. ABSTRACT
          The report gives results of a laboratory study of the control of SO2 emis-
  sions from copper smelters by H2S production from copper concentrates. Digestion
  of neutral roasted pyritic copper concentrates with HC1 was studied as a means of
  producing sufficient H2S  for reaction with the SO2 from converter gas to produce
  elemental sulfur and then minimize SO2 emissions from copper smelters. In this
  step the copper sulfides are maintained in insoluble  form.  A large fraction of the
  iron was shown to be solubilized with equivalent production of concentrated H2S at
  relatively low temperatures. In such a process a pure form of iron oxide would be
  produced as a by-product during acid regeneration.  In the  integrated scheme, diges-
  tion would likely be the only step which is not a state-of-the-art process. An econ-
  omical overall process appears probable to yield elemental sulfur and also pure iron
  oxide as useful by-products. Costly reverberatory furnaces would be eliminated.
  Needs for slag forming minerals would be greatly reduced. With efficient utilization
  of excess pyrite, increased recovery of copper and precious metals from the ore
  body should also be realized. Further laboratory studies are recommended in pre-
  paration for pilot plant investigations.
                             KEY WORDS AND DOCUMENT ANALYSIS
                 DESCRIPTORS
                                          b.IDENTIFIERS/OPEN ENDED TERMS
                                               c. COSATl Field/Group
 Air Pollution
 Sulfur Dioxide
 Copper Ores
 Smelters
 Smelting
 Hydrogen Sulfide
Hydrochloric Acid
Air Pollution Control
Stationary Sources
Pyritic Copper
13B
07B

11F
13H
 8. DISTRIBUTION STATEMENT
                                           19. SECURITY CLASS (This Report)
                                           Unclassified
                                               21. NO. OF PAGES
                                                    117
  Unlimited
                       20. SECURITY CLASS (Thispage)
                       Unclassified
                                               22. PRICE
EPA Form 2220-1 (9-73)

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