Copper Industry

2.3 COPPER INDUSTRY

The copper industry consists of all establishments in the following SIC categories:

SIC 3331 Primary Smelting and Refining of Copper - Establishments primarily engaged in smelting copper from the ore, and in refining copper by electrolytic or other processes.

SIC 33412 Secondary Copper - Establishments primarily engaged in recovering copper from copper and copper- based alloy scrap by utilizing a variety of melting and refining methods.

SIC 3351 Rolling, Drawing, and Extruding of Copper - Establishments primarily engaged in rolling, drawing, and extruding copper, brass, bronze, and other' copper- base alloy basic shapes, such as plate, sheet, strip, bar, and tubing.

SIC 3362 Brass, Bronze, Copper, Copper Base Alloy Foundries (castings) - Establishments primarily engaged in manufacturing castings and die castings of copper and copper-base alloy.

The domestic copper industry is segmented into primary and secondary sectors on the basis of whether the copper product has originated from mined copper (virgin ore) or from scrap. By this definition, firms in the primary sector preddinantly transform mined copper into refined copper; firms in the secondary sector either predominantly process scrap into secondary refined copper or prepare it for direct consumption in the form of unrefined copper scrap. The following sections detail these two divisions of the copper industry.

2.3.1 Primary Copper

2.3.1.1 Industry Structure. In 1979, nine companies operated 17 primary smelters with a smelting capacity of approximately 8.2 million tons of charge estimated to represent i.4 miilion tons of smelter product. Re- finery capacity totaled 2.6 million tons, of which 88 percent was electrolytic refining and electrowinning capacity and 12 percent was fire-refining capacity (6) . According to the Bureau of Mines, 4,707,916 tons of copper was produced in mines, smelters, and refineries from domestic and foreign ores in 1979 (5).

Mills are almost always located close to the mines to minimize transportation costs. The value of the concentrates is high enough to allow some flexibility in smelter location. With the major copper mines centered in the western states, most of the smelting capacity is in that region .

The copper industry is highly concentrated. In 1979, four companies accounted for 63 percent of the domestic mine production. Twenty-five mines accounted for 94 percent of the U.S. output, the 5 largest produced 45 percent (6).

2.3.1.2 Process Description.* The three major steps in producing copper metal are roasting, smelting, and refining. Each is described below.

Roasting. The object of roasting copper s fide ores and concentrates is to regulate the amount of sulfur so that the material can be efficiently I smelted and to remove certain volatile impurities such as antimony, arsenic, - and bismuth.

Roaster gases are rich in so23 such that sulfuric acid can be produced in large quantities; some of the acid is used in the adjacent electrolytic refinery. Some copper smelters do not roast prior to reverberatory furance smelting but do make sulfuric acid from converter gases; some smelters do not have adjacent electrolytic copper refineries. Acid plant blowdown slurry effluent .yields a sludge which is not 100 .percent recycled, as will be discussed in a following section.

Smelting. Roasted and unroasted materials are smelted after mixing with suitable fluxes in reverberatory furnaces. In this reducing or neutral atmosphere, copper and sulfur form copper sulfide and iron sulfide. The combination of the two sulfides known as copper matte collects in the lower area of the furnace and is removed. Mattes comonly contain 40 to 45 percent copper. Impurities such as sulfur, antimony, arsenic, iron, and precious metals are in the matte. The remainder of the molten mass containing most of the other impurities floats on top of the matte and is drawn off and discarded as a slag.

A simplified flow diagram identifyfng solid waste sources and disposal is shown in Figure 2-4. The primary process steps comprise (1) roasting to reduce sulfur content, (2) reverberatory furnace smelting to form copper sulfide matte and a siliceous slag which is discarded, (3) oxidation

* This section was derived from Reference 2.

38 i Copper Sulfide LIine Rock Silica Reverts Oust to

ACID PLANT

blowdarn Slurry

Scrubber Slurries

I41 scel lanrous

Overflow

To Tailings Pond (0lrsolv.d folldr Content 280 lbr. wr Short Ton Sopwr Productj

Ann* capper Product

Figure 2-4. PRIMARY COPPER SMELTING AND FIRE REFINING

Source: Reference .2 (blowing) the molten sulfide matte to form molten “blister” copper and iron silicate slag which is returned to the reverberatory, and (4) furnace purification (fire-refining) of the molten copper such that anodes suitable for electrolytic refining can be cast (or copper product that can be marketed directly).

Converting is the final stage in the process of smelting and is accomplished by blowing thin streams of air through the molten matte in a refractory-lined converter to oxidize the ferrous sulfide to sulfur dioxide, to eliminate the sulfur as a gas, and also to form a ferrous slag containing trace metal impurities. When the converter cycle is finished, the converter is tilted to discharge the lighter slag and then the relatively pure copper metal, referred to as blister copper, into ladles in which it is transferred to a holding furnace and then to a poling furnace for fire-refining or casting into anodes for electrolysis.

The blister copper produced by smelting is too impure for most applications and requires refining before use. It may contain silver and gold, and other elements such as arsenic, antimony, bismuth, lead, seleuium, tellurium, and iron. Two methods are used for refining copper--fire refining and electrolysis.

Fire Refining. The fire-refining process employs oxidation, fluxing, and reduction. The molten metal is agitated with compressed air. Sulfur dioxide is liberated and some of the impurities form metallic oxides which com- bine with added silica to form slag. Sulfur, zinc, tin, and iron are almost entirely eliminated by oxidation. Lead, arsenic, and antimony can be removed by fluxing and skimming as a dross. Copper oxide in the melt is reduced to metal by inserting green wood poles below the bath surface (poling). Re- ducing gases, formed by ‘combustion of the pole convert the copper oxide in I:.. the bath to copper. If the original material does not contain sufficient gold or silver to warrant its recovery, or if a special purpose silver- Containing copper is desired, the fire refined copper is cast directly into forms for industrial use. If it is of such a nature as to warrant the recovery of the precious metals, the fire refining is not carried to completion but only far enough to insure homogeneous anodes for subsequent electrolytic refining.

Electrolytic Refining. In the electrolytic refining step anodes and cathodes (thin copper starting sheets) are hung alternately in concrete electrolytic cells containing the electrolyte which is essentially a solution of copper sulfate and sulfuric acid. When current is applied, copper is dissolved from the anode and an equivalent amount of copper plates out of solution on the cathode. Such impurities as gold, silver, platinum-group metals, and the selenides and tellurides of metal fall to the bottom of the i tank and form anode slime or mud. Arsenic, antimony, bismuth, and nickel enter the electrolyte. After the plating cycle is finished, the cathodes are removed from the tanks, melted, and cast into commercial refinery shapes. The copper produced has a minimum purity of 99.9 percent. Important auxiliary process steps are slimes recovery from the cell bottoms, and electrolyte purification to permit electrolyte reuse and to recover materials of value. The slimes tend to be rich in Se, Te, As, Ag, Au, and Pt such that their value is very high; the typical refinery ships the slimes filter cakes elsewhere for treatment to recover these metals. (Three of the existing electrolytic copper refineries in the U.S. treat slimes from their own operations and from other refineries.) Since slimes yield is only of the order of 0.002 short'ton pershort ton of copper refined, it is ap- parent that the slimes recovery "industry" in the U.S. is very small relative to copper production.

Electrolytic purification consists of the following steps performed on the bleed stream of impure sulfuric acid containing copper and other elements in solution: (1) copper removal by electrolysis with insoluble lead anodes in liberator cells; (2) filtration to remove an arsenical sludge; and (3) evaporation to precipitate nickel sulfate. A "black acid" product remains after nickel sulfate removal. This is marketed for acid recovery.

Slags taken from the melting and refining furnace (and from the anode casting furnace at refineries wich cast anodes) are very rich in copper (50 percent Cu), i.e., immediate recycle to a copper smelter is dictated and practiced.

2.3.1.3 Waste Stream Characteristics.* ~e sources of solid waste at copper smelters include furnace slag, wet sludges, and in a few cases, collected particulates. Solid wastes at electrolytic copper refineries consist of only a few miscellaneous sludges and slurries; these occur in relatively small quantities because of the high purity of the material fed to the refinery process.

Slag is the major solid waste from copper smelting. It is usually tapped from the furnace, transported molten to the slag pile, hot dumped, and allowed to solidify. At some copper smelters the slag is granulated as it is tapped from the furnace. Particulates are collected from the off-gases and are usually recycled to the process, although those with accumulations of impurities may need to be discarded. The furnace off-gas also contains sulfur dioxide.

* This section was derived from Reference 128.

41 Slurries containing suspended and dissolved solids arise from a number of Sources including acid plant blowdown, DMA purge, wet scrubbers, blister cake cooling water, anode cooling water, other contact cooling water, and plant washdown water. At some plants the slurries are discharged to a small lagoon for settling, then the solids are dredged and recycled for metal recovery. At these plants this material does not constitute a solid waste. At other plants the slurries are either discharged directly to a tailings pond (when ore concentrating operations are nearby) or to a lagoon, where solids settle and remain, or are periodically dredged and disposed of on site. At these plants the slurries constitute a solid waste.

At all primary copper smelters the collected particulates contain high concentrations of recoverable metals. The preferred practice is to immediately recycle this residual to the process. At a few plants the particulates are sold to lead smelters after several cycles through the copper smelter, and at one western smelter the dusts from the roaster and reverber- I atory flues are processed on site in an arsenic plant to produce arsenic 1 trioxide. By no means do these materials constitute a solid waste. At a few plants the collected particulates are temporarily stored on site before recycling, but this involves a small amount of material and the storage time is generally short.

The following table summarizes the amounts of wastes generated by the primary copper industry in 1980. These amounts include wastes generated by all the different smelting processes employed by the copper industry. 1'. Table 2-1 SOLID WASTE FROM THE PRIMARY COPPER SMELTING AND REFINING INDUSTRY IN 1980

Amount Waste (tons per year)

Slag 3,271,500 Acid Plant Blowdown .. ;12,000 Miscellaneous Slurries 27,000 Refinery Slurries -3,600 Total 3,314,100

Source: Reference 128.

! The characteristics of furnace slag vary with the concentrate feed material and with furnace operation. The grade of matte produced affects the furnace slag, the higher grades of matte yielding a higher percentage of copper in the slag. Both Noranda and flash furnaces produce a high grade of matte, and both processes incorporate secondary slag treatment to recover the copper. Typical furnace slag analysis appears in Table 2-2.

Table 2-2

TYPICAL EURNACE STAG ANALYSIS--PRIMARY COPPER -1/

Constituent Percentage

cu 0.75 . Zn 0.40 Mn 0.028 Sb 0.025 Pb 0.014 Cr 0.011 Se 0.003 Ni 0.002 Cd Trace k8 Trace

-1/ Values adapted from Reference 2, page 19.

Characteristics of acid plant blowdown fluctuate with the composition of the ore concentrate and with the gas cleaning operation upstream of the acid pl . Acid plant blowdown is acidic and usually contains copper, lead, zin iron, and arsenic, among other components. Analysis of acid plant blowdm from th s appears in Table 2-3.-

Miscellaneous slu the smelter and refinery originate , from a variety of sources a racteristics vary widely. Slurries from the smelter probably would contain copper, lead, zinc, mony, and bismuth; slurries from refinery would contain these,elements pius sulfuric acfd, copper sulfat nd any impurities indigenous to the ore. Data were not available for characterization of these slurries. I Table 2-3 TYPICAL ACID PLANT BLOWDOWN WASTE ANALYSIS (pounds per ton except as noted)

Parameter

PH 2.0 so 4 72.0 HCN 0.0016 As 0.118 Cd 0.0194 cu 0.0020 Fe 0.0464 Pb 0.1796 HI3 0.0002 Ni 0.0020 Se 0.0180 Te 0.0000 zn 0.436 Oil and grease 0.0

Source: Reference 129.

2.3.1.4 Potential Hazardous Wastes.

Three general types of solid residues containing potentially hazardous elements are found at copper smelters, namely slag, dry dusts, and wet sludges and slurries. Not all copper smelting and refining wastes are considered as having equal degrees of potential hazard (2).

1. Copper Smelting Dusts and Sludges

In solubility tests described in Reference 2, it was found that sludges (acid plant sludge) and dusts (converter dust, reverberatory dust) leached copper, lead, zinc, and cadmium in significant concentrations. These wastes are therefore all considered potentially hazardous (2).

Acid blowdm sludges contain the following concentrations of elements : . ..

,.. 2.3 2 Secondary Copper

2.3.2.1 Industry Structure.* Recovery of copper from scrap is not conducted entirely at secondary copper smelters. Of the 1.5 million tons recovered in 1979, 20 percent was from primary producers, 36 percent from brass mills, 39 percent from secondary smelters, and the remaining 5 percent from chemical plants, foundries, and manufacturers (5).

In the secondary copper smelting and refining industry in the United States, there are approximately 50 producers of either brass and bronze ingots or secondary refined copper operating approximately 70 plants. The plants in this industry fall into two fairly distinct segments: (1) producers of brass and bronze ingots; and (2) producers of unalloyed copper.

Most of the firms in the secondary copper industry are small, indi- vidually owned operations having only one plant; only a few are publicly held. A minority of the firms (but still representing a large fraction of the production) are either subsidiary operations of large mining companies or are sub- sidiaries of conglomerates.

There are 63 plants in the brass and bronze ingot segment. These producers manufacture a wide variety of specification alloys. These alloys . generally*fit a series of specifications that have been outlined by both ASTM and the Brass and Bronze Ingot Institute (BBII) and are in the form of 30 pound brass or bronze ingots. Some of the smelters also produce a series of materials in the form of shot that are sold to factories for the inoculation of gray iron. The shot may be pure copper or copper-nickel alloys of various types.

There are seven producers of unalloyed copper, which may be in the form of blister copper, fire-refined copper, cathode copper, wire bar, continuous cast, or a finished product, depending on both the production scheme and the needs of the customer. Also, several precious metals are usually recovered as a result of electrorefining to produce cathode copper.

Producers of brass and bronze ingots are located mostly in the northeastern, Pacific coast, and east north central states. Producers of unalloyed copper are also located in heavily industrialized areas, mostly in the northeastern states, with one plant in the south and two in Illinois.

* This section was derived from Reference 1.

4 46 2.3.2.2 Process Description.* The basic raw material of the secondary copper industry is copper and copper-based alloy scrap. About two-thirds of the amount of secondary copper recovered is in the form of either brass or bronze, while one-third is in the form of unalloyed copper.

The blast furnace is used extensively in secondary smelters for smelting low-grade copper and brass scraps, refinery slags, drosses, and skimmings to produce black copper (80 to 90 percent Cu). The conventional secondary copper blast furnace is a top-charged, bottom-tapped shaft furnace heated by coke burning in a blast of air introduced through tuyeres placed symmetrically around the bottom of the shaft. The black copper produced in the blast furnace is converted to blister copper by blowing air into the molten charge.

About two-thirds of the secondary copper production in the United States is used in ingot plants and foundries to make brass and bronze alloys by simple melting and refining methods. Melting, refining, and alloying procedures are essentially the same r reverberatory, rotary or crucible furnaces.

Capacities of stationary reverbe ory I furnaces used in secondary smelters range from a few thousand pounds to 100 short tons or more. The side- or end- charged arched-roof tapping furnace is the type most extensively used. Rever- beratory furnace slags usually contain metal values that can be recovered in the blast furnace. Slags produced by-small secondary plants are frequently sold to primary smelters on the basis of copper content only. Some plants grind the slag and recover met lic constituents in milling operations before the slag is sold.

The rotary furnace is designed to provide efficient melting and refining and convenient pouring of.fairly large melts. Furnace capacity ranges from several short tons to 50 or more short tons of nonferrous metals. It may have a *particular advantage over stationary furnaces for melting loose or baled light scrap, because the rotary mixing action promotes better heat transfer to the melt and causes a more rapid coalescence of melted globules.

A fairly large tonnage of s ondary copper is produced in crucible furnaces, heated by gas, oil, coke or electricity. The once popular coke- fired pit furnace is seldom used today. Crucible furnaces are used in the secondary-copper industry for melting clean, well-segregated scrap- in foundries. Very little fire-refining is performed in crucibles.

?leltisg furnaces are always associated with other equipment designed to receive the melt. Melts are usually tapped from furnaces into feeder ladles, which transport the metal to a mold line for making conventional ingots. The . mold line is a series of ingot molds placed on a rack that may be stationary or movable. If stationary, the molds are filled with metal poured from a portable ladle.

* This section was derived from Reference 1.

47 . I

The electrolytic refining of secondary copper, the copper from the I reverberatory furnace is further refined by first dissolving in sulfuric and then electrolytically redepositing pure copper on the cathodes of electrolytic cells.

2.3.2.3 Waste Stream Characterization.* The typical blast furnace dust will have the following chemical composition:

Zn 58-61 percent Pb 2-8 percent Sn 5-15 percent cu 0.5 percent Sb 0.1 percent i c1 0.1-0.5 percent

Because of the high zinc content, the dust can be recycled for zinc compound manufacture largely as point additives. This waste by-product is thus sold to chemical and plant manufacturers. i The typical blast furnace slag will have the following composition:

FeO 15-20 percent CaO 7-11 percent i si0 30-34 percent 4-6 percent 43cu 1.6 percent I . Sn 0.7 percent Pb 0.3 percent I Zn 1.0 percent A typical plant will produce about 3,800 tons of blast furnace slag annually. This waste is generally disposed of by either on-site or off-site iI .open duping. The material is present mainly in a silicate type chemical structure with high density, hardness, and impact strength.

. In electrolytic refining, lime treatment of spent electrolyte I generates a predominantly lime sludge containing significant concentrations of nickel, zinc, copper, chromium, and cadmium. This sludge is generated at a rate of approximately 3.3 pounds wet sludge per ton of copper. 1I

! I

1I I * This section was derived from Reference 2. I

.rp 48 2.3.2.4 Potential Hazardous Waste Streams

1. Blast Furnace Slag

The typical blast furnace slag will have the following composition: FeO, 15 to 20 percent; CaO, 7 to 11 percent; Si02, 30 to 34 percent; Al2O3, 4 to 6 percent; Cu, 1.6 percent; Sn, 0.7 percent; Pb, 0.3 percent; Zn, 1 percent (2). Although this material is dense and hard, solubility tests showed significant concentrations of soluble zinc, cadmium, copper, and lead. This waste slag is therefore considered potentially hazardous (121).

'Solubility Tests Filtrate Data in mg/l

Zinc 55 Lead 6 Cadmium 1.0 Antimony <0.2 Chromium 0.03 Tin

Source: Reference 121.

i c The Soderfors process (188) as it is so named, utilizes the capacity of undissociated acids such as NH03 and HF to form adduct complexes with certain types of organic compounds, for example, tributylphosphate (TBP). Since the adduct complexes are soluble in, for example, kerosene, the acids can be extracted from an aqueous solution by an organic solvent.

The basic flow scheme in the Soderfors process is: (a) addition of H2SO4, and the subsequent extraction of NH03, HF, and Mo (probably as molybdic acid) by means of an organic solvent consisting of 75 percent tributylphosphate dissolved in kerosene, (b) stripping of HNO3 and HF from the organic solvent with water and recycling of the acids to the pickling bath, (c) precipitation of the metals Cr, Fe, and Ni from the raffinate by means of NaOH; CaC03 is added to form CaFe and to act as filter aid, (d) scrubbing of the molybdenum from the organic solvent, which is then returend for further extraction. ,

A full scale plant has been built at the Soderfors steel works and is designed for the treatment of pickling baths from the pickling of 70,000 tons of stainless steel annually.

4.1.5 Foundry Sand Recycling

. Sand is seemingly among the most abundant materials on earth, yet like many other things that used to be inexhaustable, it is beginning to have an increasing value. Sand is not just sand to the foundryman but a specialized material that must be at the right place at the right time. This degree of specialty, plus material handling,. the thing factor and the cost to discard it are factors that give foundry sand a meaning and a cost.

Mold cores used in foundries in the manufacture of iron castings are made of silica and bonded with organic oils and resins. During the casting process a portion of the organic binder is oxidized or decomposed and the remainder adheres to the sand. The used core sand is unsuitable for reuse due to the detrimental effect of this residual binder on sand properties.

With most foundries, once sand has been used, it is either discarded or only partially reclaimed by either pneumatic scrubbing or washing with water. However, there are several alternatives to simply disposing of spent sand, alternatives that should be examined not only because of economics but also because of increasing regulatory requirements by pollution control agencies (77j.

However, economically speaking, disposal of waste sand is very costly because the costs of replacement keeps escalating as well as the cost of delivering it. Not only is "once-through" sand costly to replace but new sand offers certain other cost disadvantages. Binder consumption is increased with new sand, for one thing, and if sand is used only once, without reclamation, valuable metal is probably being wasted that would otherwise be recovered during the sand reclamation process (77).

The foundries that partially reclaim sand do so to eliminate undesirable fines fraction which is produced in the molding line and shake- out. Partial reconditioning of the surfaces of the acceptable grains can be done in several ways (77).

Air Classification--Sand is aerated to remove unwanted fines. Although this reduces the moisture and amount of binder needed in reuse, it does nothing to restore the surface condition of the old sand. Conse- quently, reclaimed sand produced by this method does not have the versa- tility of new sand.

Dry Abrasion, Dry Classification--Lumps of sand are crushed to grain size in a sand muller. The sand is then aerated after mulling to remove the fines and metal. For this system to produce an acceptable reclaimed sand, small batches must be mulled for long periods to provide sufficient abrasion to produce a good sand grain.

Pneumatic--Used sand grains are introduced into an air stream and blasted at a fixed target to loosen and peel of€'the grain coatings. The same air stream conveys the fine material to a bag collector. The sand grains are returned for use and the debris is disposed of by truck and dumped.

Roto Conditioning--Mechanical spinners pump the sand at a high velocity in the dry stage. The turbulent flow causes one grain to grind against another, producing the scrubbing action that removes used binder coatings.

Wet Scrubbing, Classification, and Drying--First, the used sand is mixed with liquid to produce a slurry. The slurry moves to agitators for a preset detention time. Agitation produces an intense scouring action which removes clays, char, and other binder deposits as well as breaklng up clumps. The agitator then discharges the slurry to a classifier where desliming takes place to remove contaminating fines and other binder debris. After desliming, the reclaimed sand is dewatered in a horizontal filter or a vertical drum filter and dried in a fluid bed or rotary dryer. Feed is continuously metered and introduced above the bed where hot gases classify and dry the sand. The fine product, thoroughly dried, is carried aloft with the gas and is recovered in the dust collection system. Coarse product flows across the construction plate and is discharged for return to the foundry for use. All these methods have been used primarily on clay-bonded sand. However, there are more complex binder and chemical resins used in many foundries today, and scrubbing alone cannot provide total sand reclamation. With chemically bonded sands, such as furan, or ureaformaldehydelfurfuryl alcohol, or such catalysts as phosphoric acid or taratoluene sulfonic acid to name but two, thermal treatment has been shown to loosen the adhering debris from the grains or separating small amounts of residual impurities from the bulk of the sand (77).

Thermal reclamation is achieved in a fluidized bed calciner which consists of a calcining bed and one or more heat exchange beds for heat recovery.

The sand is metered from the feed bin and introduced into the top (preheating) bed of the reactor. Here it is dried and heated by hot gases from the calcining compartment. The fine fraction entrained in the gas is carried over to a cyclone where it is collected and discharged to the calcining bed. Size separation takes place with coarse sand passing to cooling compartments where it serves to heat incoming air before discharging to the fluid bed aftercooler. Heated gas flows through the system in one direction while the solids flow in the opposite direction, thereby accomplishing counter- current heat exchange. The fine fraction from the calcining compartment is collected in cyclones and is discharged either to the aftercooler where temperature is reduced to approximately 100°F for reuse in the foundry, or to a dry scrubber if needed (77).

For many years the Japanese have recognized the need for thermal reclamation because of the high cost of their new sand. Currently there are nine fluid bed thermal reclamation.units in operation in Japan (77).

For additional reading see References 77, 78, 89, 102, 156, 248, 249, 250, 251, 252, 253.

Additional research on utilization of foundry wastes can be found in References 79 and 80. In summary it was found that the use of arc furnace dust for noncorrosive primer paint pigment constitutes the most promising use for any foundry effluent. In addition, arc furnace dust might be used as tint and tinting hues in furniture stains and ink. Fine particle size, compati- bility, hiding power, and tinting strength govern these applications. Sec- ondary applications for arc furnace dusts are road fill and ballast, and for road safety applications.

Cupola effluent, CD, was suggested as a soil conditioner due to its granular nature. However, more research into the effect of this use is sug- ges ted. Sand reclaimer waste, due to its gel-like nature when wet, may be useful in stabilizing soils to prevent collapse during evacuation. In addition, a mixture of this waste and Portland cement or silicates injected into the soil may assist in leak prevention in foundations.

4.2 ALUMINUM INDUSTRY

The major process waste in production of alumina is the leach residue from the caustic digestion of bauxite, red mud. This waste mud is im- pounded to prevent the contamination of surface waters. In the United States, over 10 million tons of mud waste is generated annually. This mud is im- pounded in lakes built adjacent to the refineries, ranging in size of 2,000 to 3,000 acres. Mud impoundment is not an ideal solution to the disposal problem. The dikes of mud lakes must be maintained, and there is always the risk of a break and spill of the mud into a nearby stream or waterway.

This section deals with the current methods and proposals for utilizing the red muds. The utilization methods have been divided into three categories: Metals Recovery, Construction Products, and Miscellaneous.

Metals Recovery

Red mud is essentially a low-grade iron ore, containing typically 30 to 60 percent iron oxides. Recovering this iron and/or other metals would utilize an available high tonnage waste material. Several processes exist to recover metals from red mud, but it is not believedtbtany of these are in commercial operation..

Orre process, developed by McDowell-Wellman Engineering Company (now owned by Dravo Corporation), converts red mud into steel billets (44, 18).

In this process, the red mud, either in the form of a paste-like, low-moisture (20 percent) cake or a dry powder, is pelletized with coke into balls of about 1/2 inch in diameter. These pellets are charged into a dry- - - ing and prereduction unit. A layer of limestone as a flux is distributed on top of the bed of pellets. The bed is fired to a temperature in excess Of 2,300°F, calcining the limestone to lime and prereducing the iron oxide -in the pellets. Carbon monoxide from the partial combustion of carbon contained in the coke forms carbon dioxide with oxygen from the iron oxides in the red mud. At this stage, the pellets are chemiclly reduced, having lost 50 to 70 percent of their oxygen content.

95 The hot pellets are next fed into an electric furnace where smelting takes place at about 2,900"F. This furnace provides the energy to complete the reduction reactions between carbon, carbon monoxide, and iron oxides to form molten iron. A slag floats to the top, the result of reactions between the lime flux and the silica, caustic, alumina, and titanium dioxide found in the red mud feed. Certain red muds may contain quantities of phosphorus, and if the coke is high in metals such as vanadium, these elements will be found in the molten iron.

Refining and conversion to steel is performed in a Basic Oxygen Furnace (BOF). Vanadium, titanium, and silicon are slagged off when oxygen is initially injected through the lance. The next component to be oxidized in the BOF is carbon. After the carbon has been partially removed by the oxygen, a slag containing oxides of vanadium, titanium, and silicon is poured off. Lime is then added for a final oxygen-blowing cycle. The lime removes the phosphorus as a new molten slag which is recovered, cooled, and granulated. This by-product may be sold as a fertilizer.

Molten steel from the BOF is fed to a casting unit which molds the liquid into steel billets.

This process was performed in a pilot plant in Cleveland, Ohio, and operated successfully for over eight years. The plant has since been partially dismantled and no longer operates (258).

In a pilot scale research experiment performed by the Bureau of Mines (16), both the carbon-lime-soda sinter process and electric-arc smelting with coke and limestone were employed to recover iron and alumina. Al- umina recoveries of 85 to 90 percent were obtained by the sinter process; although iron recovery was about 80 percent, the magnetic product was finely divided and'difficult to separate from the gangue. Smelting red muds in the electric-arc furnace produced a pig iron with 98 percent recovery of iron; up to 84 percent of the A1203 was recovered by leaching the calcium aluminate slag with Na2CO3 solution. Also, a titania product assaying 96 percent Ti02 was recovered from the nonmagnetic portion of Surinam red mud sinter by extraction with H2SO4.

In a Yugoslavian sponsored investigation (25, 26) large scale smelting of red mud was conducted, in blast and electric furnaces, to produce an alloy or high-quality pig iron and slag, The briquetted, pelletized or sintered red mud had to be smelted in such a way that alumina and titania went into slag, as well as all minor and rare-earth elements, while the alloy elements were mostly reduced to pig iron. 1. The granulated slag, formed during the smelting Process, is leached in a suitable tank where it is vigorously agitated With 30 Percent sulfuric acid, for 30 to 60 minutes, at the temperature of 80 to 90°C- After the leaching, undissolved slag is separated by decanting. The undissolved slag residue, i without any further washing out or drying, was mixed with an appropriate amount of raw phosphates in order to produce a new type of Synthetic fertilizer, well known in Yugoslavia, Czechoslovakia, and India under the name of PELOFOS.

From the clear liquor obtained by filtering or decanting of the undissolved slag, compounds of Ti, Zr, U, Th, and lanthanides are extracted by di-a-ethylhexyl-phosphoric acid and a-ethyl-hexanol dissolved in kerosene. From the solution, after extraction of titanium and other elements mentioned above, iron compounds and aluminum sulphate or alumina are recovered.

In a Czechoslovakian report (15) details of an experiment for vanadium recovery from red mud are given. The sintered red mud is charged into a blast furnace where it is reduced. The pig iron, containing up to 1 percent vanadium oxide (five times as much as the charged sinter) is oxidized in the ladle to yield vanadium-rich slag. The V2O5 that forms a fairly high proportion of this slag can then be further reduced into metallic vanadium or ferrovanadium by any of the usual methods.

The U.S. aluminum industry has accumulated approximately 85 million tons (dry basis) of Bayer Process mud wastes at ten localities. Although the uranium content is low (10-40 ppm U308), the muds are estimated to contain 1,874 acid-extractable tons of U308. Total extractable thorium is estimated to be 11,009 tons (254).

In a recent study (254) it was found that laboratory acid extraction leaches 97 percent of the total u308 from some types of muds, but recovers less than 50 percent U3O8 in others.

If the anticipated complex extraction problems can be overcome, Bayer Process muds could be assigned a place among other potentially recoverable low-grade uranium and thorium resources.

Construktion Products

For most of the possible uses of red mud in construction products, it would be necessary to dry or at least partially dry the mud, for example, before firing inabrick kiln. The cost of drying or dewatering the red mud will often make the use of this waste material economically unattractive over conventional materials. However, cement, building blocks, bricks, and aggregate are applications where red mud might be used.

97 In one investigation it was shown that a suitable heat-treatment followed by ball milling to cement fineness, a material with pozzolanic ac- tivity, referred to as bauxite pozzolana, can be produced (31). Without sacrificing acceptable compressive strengths, up to 40 percent replacement of Portland cement by bauxite pozzolana seems possible in the preparation of concrete. A 20 percent replacement of Portland cement with bauxite pozzolana gives concrete of a higher compressive strength. The increase in strength has been attributed to the formation of a calcium-aluminate hydrate bond in addition to the calcium-silicate-hydrate bond of Portland cement concrete.

In another cement product application, pressure filtration of red mud is used to reduce moisture and produce a cake which can be handled by conventional soil handling equipment (23). The cake has been shipped to a cement company for use as a raw material (Fe2O3) of cement production. Compared with conventional ferric oxide materials for cement production bauxite residue is acceptable for cement production and is expected to be used at other cement plants in Japan, where this report originated.

Synthetic dense aggregates have been produced from red mud in the U.S. and Japan. In both cases the red mud was pelletized and fired, in the U.S. work at temperatures of 1,260 to 1,316"C (12) in the Japanese work at 1,200"C. The Japanese aggregate is reported to have a saturated surface. dry relative density of 2.67 and a water absorption of 1.5 percent. Its properties in concrete have been tested in comparison with a river gravel and the compressive strength, bending strength, and tensile Strength of the concrete made with the red mud aggregate have been found'to be superior (27)

Methods were developed by researchers at IIT Research Institute, Chicago, to,use red mud wastes for the production of lightweight building materials. A mechanical foaming process produced strong, uniform structures. Densities ranged from 30 to 70 pounds per cubic foot and compressive strengths from 150 to 2,200 psi (30). By controlled processing, foamed block could be formed with properties varying from insulative to structural.

The use of red mud for making structural bricks and blocks has been considered in numerous foreign publications, particularly German and Japanese. Unfortunately it was impossible to have these translated and therefore much research is being overlooked.

However, as reported in Reference 34, it was fccnd that the resulting bricks and blocks were mechanically weak. The compressive strength of the standard U.S. common brick exceeds 8,000 psi. In most cases, the red mud products had strengths less than this. This was true even in the case of brick made of mixtures of red mud and clay fired at temperatures comparable to those used in firing all-clay brick (256, 257). A 5 to 10 percent addition of red mud to clay was used to make roofing tile (21).

I It is judged from the literature that it is not likely that the red mud content of brick can exceed 50 percent and that 25 percent is a more likely prospect. Brick with more than 50 percent red mud were mechanically weak by U.S. standards (34).

Miscellaneous

Red mud from the Bayer Process contains sodium as sodalite. Many investigations have been made on a method using sodalite as an SO2 absorbent, and the Sumitomo Aluminum Smelting Co. of Niihama, Japan has developed a process for the SO2 removal from the waste gas stream generated in a power plant (46).

The RMDS process, as it is referred to, is being used at the power plant near the company's Kikumoto Works, where alumina is produced. The claimed efficiency of the SO2 removal is about 96 percent (46).

A U.S.-patent No. 4,071,373 (8) describes a process for the manufacture of aluminous cement from aluminum smelting residues. The steps consist of preparing a mixture of 150 parts of aluminum smelting residue containing at least one of aluminum carbide and aluminum nitride. 50 parts of limestone, 50 parts of blast furnace slag, 50 parts of converter slag, and 30 parts of aluminum hydrosilicate. The mixture is then crushed.and sintered to form a clinker. Powdering the clinker forms the cement.

In another patent, U.S. No. 4,179,279 (20) a process for treating molten steel-slag with red mud to improve the structural stability of the slag and to render it suitable for landfill and construction material as well as use as a road base rqaterial, is described.

An interesting possible use of red mud is as a coagulant for the treatment and. sedimentation of municipal wastewaters and for sludge floc-' culation. This has been done on a commercial scale in England before the war. The process involves treatment of the red mud with concentrated acids, then drying at modest temperatures, and crushing to give a granular product. Either..sulfuric or hydrochloric acids can be used to convert iron .and-aluminum oxides in the red mud to sulfates or chlorides. Thus, ferric chloride or alum can be produced which are well known for their coagulating properties. In addition to the effects, it has also been reported that'the red mud materials are effective for removing phosphorus from the municipal wastewaters with 70 to 80 percent effectiveness (34).

It was also reported that red.mud as a fiiler in rubber gave a rup- ture strength higher than all other fillersexcepting carbon black (34). It should thus be a suitable substitute for carbon black in some applications. This use as a filler requires that 'the red mud slurry be acid washed, dried, and powedered.

99 4.2.2 Furnace Dross Recycling

During the smelting of aluminum, there forms on the surface of the molten metal a skim or dross consisting of aluminum oxide, aluminum nitride, and various other impurities. This dross contains a substantial amount of entrained metal, often 25 to 80 percent by weight of the dross, which repre- sents a significant metal loss if discarded. Efforts are made to recover as much of the aluminum metal as possible. No one method of treatment is suitable for all types of dross.

One commercial process of aluminum recovery involves crushing, sizing, and smelting of dross under a salt flux, which consists of an eutec- tic solution of sodium chloride and potassium chloride plus 1 to 5 percent cryolite. A mixture of dross and flux in the ratio of 1 part oxide to 1 part flux is heated in a rotating furnace to a temperature above the melting point of aluminum and the flux. The impurities remain solid. The molten flux selectively wets the impurities of the dross and promotes separation of the molten metallic aluminum from the solid impurity fraction. At the same time, the aluminum is protected from oxidation. After smelting is completed, molten aluminum is tapped from the furnace; the residue, also called a high- salt slag, remains (28). Previously these salt slags were dumped in landfills but due to problems of the salt leaching and contaminating surface and ground waters, this practice is unsatisfactory (28). Recycling of these slags will be recovered later in this section.

A Bureau of Mines reports describes a salt-free, fluxless method for reclaiming metallic aluminum from aluminum dross (28). The dross is smelted .) under argon, nitrogen, or carbon dioxide, instead of a salt flux, in an ex- ternally heated pot furnace. Alumi.num recovery, using argon, is equal to recoveries reported from commercial smelters that process dross in rotary- type furnaces. This salt-free method could offer a reduction in air pollution, lower energy requirements by elimination of salts, and the production of up to 60 percent less solid waste for disposal than the salt method.

Another method employed in the industry is fluxing the aluminum melt with gaseous chlorine before the dross is removed. This initiates an exothermic reaction and a portion of the aluminum contained in the dross will drain out. After treatment with the exothermic fluxes, the resultant dross will often contain only 20 to 30 percent aluminum making it barely worth processing further. Low capital and operating costs are an advantage (43)

Another development has been to devise a one-step procedure using a low frequency coreiess Induction furnace for the recovery of metal from the drosss (43). This process involves the following steps:

(a) Press is skimmed from the furnace (b) The dross is then cooled as rapidly as possible

100 .-

(c) The cold dross and a suitable quantity of salt flux are charged into the induction furnace in which a molten heel of metal is already present. After some time and intensive action of the in- duced current, the metal in the dross is melted and released by the action of the salt flux and agglomerates in the heel of metal in the furnace. The non-metallic material floats on top of the molten aluminum and the two layers are separated from one another by careful decanting.

Attempts have repeatedly been made to return casting plant drosses to reduction cells in primary smelters. In this way it was hoped to recover in one step the metallic content, the alumina, and any fluorine compounds. Unfortunately, it was found that these additions seriously upset the operation of the reduction cells.

Investigation has shown that the,main upsetting compounds are aluminum carbide and aluminum nitride. It has been found that, by hydrating dross with low pressure steam, the carbi s and nitrides will decompose and the dross, after drying, can then be added to reduction cells (43).

Disposal of salt cake or salt slag residues as landfill has been practiced for many years, but this practice has been outlawed in seven states because of water contamination. Moreover, salt cake and slag react with water to form amonia and methane gas. As salts-leach away, large subsurface voids occur into which anyone falling would suffocate from these gases (40).

The Bureau of Mines has'developed a hydrometallurgical process.for treating salt slags from aluminum dross furnaces at its College Park Metal- lurgy Research Center (14). A 100-pound-per-hour minipilot plant has been constructed'to demonstrate the process. In this process, the salt slag is leached with water at room temperature. Filtering the leach solution recovers an aluminum-rich fraction, which is returned to the dross furnace, and a fine aluminum-oxide fraction. Evaporation of the filtrate recovers a high-purity fluxing salt.

4.2.3 Spent Potlining Recycling*

Used potlinings consist largely of carbon but also contain appre- ciable quantities of fluorides, aluminum carbide, and nitride. The car- bides react with water vapour in the air to produce methane, hydrogen, acetylene, and other hydrocarbons. The nitride similarly yields ammonia. These traits render the cathode material difficult and unpleasant to handle and so in the past it has been mostly dumped or used as landfill. In this section various methods of recycling are discussed.

* This section derived from Reference 39.

101 &CAN Process

This process involves the crushing and screening of the material to be processed, followed by a hydration step in which steam reacts with the car- bides and nitrides present in the material to produce methane and ammonia which are exhausted from the hydrator and burnt in an incinerator. The material is then either directly recycled or screened and recycled. The process is in operation in Kitimat, British Columbia. See also Reference 26.

Sodium Hydroxide Leaching

This process involves leaching the crushed potlinings with hot dilute caustic solutions. Cryolite is soluble in the caustic solutions and can be precipitated and recovered by treating the leach solution with carbon dioxide.

The main problem is that filter residues still contain fluorine, typically 14.9 percent F and so still present a disposal or recycling problem. See also Reference 37.

Soda Ash Fusion (17)

The Bureau of Mines conducted a laboratory investigation of a sinter-leach method to recover aluminum and fluorine from waste carbon potlining residues from secondary cryolite recovery operations in aluminum reduction plants. After removal of carbon by burning, the residue was sintered at 900" to 1,OOO"C with silica and Na2C03, then crushed and leached with dilute HaOH solution, dilute.Na2CO3 solution, water, or water containing residual Na20 which remained in the sinter after reaction and decomposition of Na2CO3. Over 90 percent of the contained aluminum and up to 87 percent of the contained fluorine were extracted from this waste product in the form of a mixed alkaline solution.

Treatment of the alkaline solution with C02 yielded a white precipitate containing 89 percent of the original aluminum and 61 percent of the original fluorine as a mixture of synthetic cryolite and alumina, together with a variable excess of soda. The precipitates assayed around 30 percent Al2O3, 13 percent F, 25 percent Na20, and up to 0.2 percent Si02.

The main problem appears to be that in the sintering mixtures, the amounts of soda ash are larger than the amounts of carbonaceous material treated.

Hot Water Hydrolysis

This approach is much the same as the ALCAN process except that hot water is used to effect the hydrolysis instead of steam. This should result

102 in a cost saving as well as preventing the production of any fluoride containing off-gases. This method appears to be perfectly feasible.

Other proposals and experimental results are presented in Reference 39. However those with negative results were not included here.

4.2.4 Foundry Slag Processing

A typical rotary furnace feed for the remelting of aluminum scrap consists of two parts scrap and one part salts. The resulting slags contain 40 to 50 percent sodium chloride, about 20 percent potassium chloride, 4 to 8 percent aluminum metal, and up to 31 percent oxides, hydroxides, silicates, and other chlorides.

In experiments conducted at a German university (10) it was found that by screening and sizing the slag, 65 percent metallic aluminum can be recovered in a concentrate of 96 percent Al.

The salts were extracted by flotation methods. The experimen to two different flotation schemes--one stage and two stage flotation.

The one stage flotation produces a chloride concentrate by indirect bulk flotation of the chlorides by use of a cation-active collector. The grade of the concentrate is better than 94 percent chlorides. NaCl recovery is about 70 percent and that of KC1 60 percent.

The two stage flotation consists of a direct flotation of sylvine and an indirect flotation of halite--both with different cation-active col- lectors. Recovery of the chlorides is about 80 percent and the concentrate grade is higher than 95 percent chlorides. The tailings, amounting to about 35 percent of the flotation feed, contain less than 20 percent chlorides.

4.3 COPPER INDUSTRY 4.3.1 Metal Recovery from Slags_- Primary Smelting Slags. The bulk of the world's output of copper is produced by smelting copper sulfide flotation concentrates in reverberating furnaces, followed by oxidizing the matte to blister copper in a converter. Slags produced in the converter are too high in copper to be sent to the dump and are returned to the reverberatory furnace for recovery of most of the copper. This method for retreating converter slag, although simple, compli- cates the operation of the reverberatory furnace and contributes significantly to loss of copper in the reverberatory slag. An economic method for retreating converter slag separately would increase furnace capacity, produce a lower-grade discard slag, and simplify furnace operations (59).

103 In experiments performed by the Bureau of Mines (59), laboratory batch flotation tests were conducted on copper converter slags to evaluate the relative merits of recovering copper from slow-cooled versus water- quenched slags. Three slags containing 1.6, 5.0, and 6.6 percent copper were used. More than 90 percent of the copper was recovered in a rougher concentrate leaving a 0.2 to 0.3 percent copper tailings when treating slow- cooled slag. Lower recoveries and higher copper tailings from 0.5 to 0.6 percent copper were obtained from quenched slag.

Cost studies showed that quenched slags can be treated at slightly lower cost than slow-cooled slags. However, the cost advantage of processing quenched slags is more than offset by the higher copper recoveries obtained from slow-cooled slags (59).

In another research project aimed at recovering copper from slags (601, a slag from a primary converter and a slag from a secondar tory furnace were leached with cyanide solutions to dissolve the copper. Under the optimum conditions, up to 90 percent of the copper could be recovered. To achieve this, a ratio of cyanide to copper of 3:1, vigorous agitation and a small amount of oxygen were found to be necessary. The rate of quenching of the slag as also noted in Reference 59, appeared to be a more important factor than the particle size.

Researchers at the Ledgemont Laboratory of the Kennecott Copper Corporation report that a two-stage pyrometallurgical process has been developed for the recovery of copper (and other non-ferrous metals) from copper smelter slag (51). All types of slags, including highly oxidized, converter-type slags, are cleaned in an electric furnace which employs mechanical mixers to promote magnetite reduction and copper extraction at . high rates. In a 10-ton pilot furnace, it has been demonstrated that slags can be cleaned to 0.2 to 0.3 percent copper with economically acceptable usage of raw’materials and energy; copper is recovered as a high-grade matte.

A description is given in Reference 69 of an integrated metallurgical process for recovery of cobalt and copper from Rokana copper ..” smelter converter slag. The slag is subjected to carbothermic reduction to produce an iron-cobalt-copper alloy suitable for hydrometallurgical treatment. The components of the alloy are separated hydrometallurgically: iron is removed as hematite during an oxidation pressure leach, copper is recovered from the leach solution by solvent extraction and electrowinning, and cobalt is recovered from the raffinate by crystallization as cobalt sulphate hexahydrate. The crystals are dissolved in recycled electrolyte and purified prior to recovery of cobalt metal by conventionai eiectrowfnning.

104 .-

I In experiments performed by the Bureau of Mines (59), laboratory batch flotation tests were conducted on copper converter slags to evaluate the relative merits of recovering copper from slow-cooled versus water- quenched slags. Three slags containing 1.6, 5.0, and 6.6 percent copper were used. More than 90 percent of the copper was recovered in a rougher concentrate leaving a 0.2 to 0.3 percent copper tailings when treating slow- cooled slag. Lower recoveries and higher copper tailings from 0.5 to 0.6 percent copper were obtained from quenched slag.

In another research project aimed at recovering copper from slags (60), a slag from a primary converter and a slag from a secondary reverberatory furnace were leached with cyanide solutions to dissolve the copper. Under the optimum conditions, up to 90 percent of the copper could be recovered. To achieve this, a ratio of cyanide to copper of 3:1, vigorous agitation and a small amount of oxygen were found to be necessary. The rate of quenching of the slag as also noted in Reference 59, appeared to be a more important factor than the particle size.

Researchers at the Ledgemont Laboratory of the Kennecott Copper Corporation report that a two-stage pyrometallurgical process has been developed for the recovery of copper (and other non-ferrous metals) from copper smelter slag (51). All types of slags, including highly oxidized, converter-type slags, are cleaned in an electric furnace which employs mechanical mixers to promote magnetite reduction and copper extraction at high rates. In a 10-ton pilot furnace, it has been demonstrated that slags can be clean'ed to 0.2 to 0.3 percent copper with economically acceptable usage of raw materials and energy; copper is recovered as a high-grade matte.

A description is given in Reference 69 of an integrated metallurgical process for recovery of cobalt and copper from Rokana copper smelter converter slag. The slag is subjected to carbothermic reduction to produce an iron-cobalt-copper alloy suitable for hydrometallurgical treatment. The components of the alloy are separated hydrometallurgically: iron is removed as hematite during an oxidation pressure leach, copper is recovered from the leach solution by solvent extraction and electrowinning, and cobalt is recovered from the raffinate by crystallization 8s cobalt sulphate hexahydrate. The crystals are dissolved in recycled electrolyte and purified prior to recovery of cobalt metal by conventional electrowinning.

104 In research work conducted by the Bureau of Mines (65), a technique was developed to recover iron from molten copper smelting furnace slags by carbon injection. Slags from gas-fired reverberatory and electric copper smelting furnaces were treated in an 800 kva electric-arc furnace by inject- ing coke breeze into the bottom of the molten bath. Over 95 percent of the iron and over 90 percent of the copper contained in the smelting furnace slags were recovered, while up to 85 percent of the carbon injected into the system was utilized. It was anticipated that the metal product containing 95 percent by weight iron and 2 percent by weight copper would be a satisfac- tory medium for recovering copper by cementation.

Secondary Smelting. In the United States, over 40 percent of the copper used by industry is obtained from scrap or secondary sources. High- grade copper scrap is generally melted in either rotary or reverberatory furnaces, and is refined by conventional methods. Low-grade secondary feed sources are usually smelted in a blast furnace.

The major waste product from the blast furnace is slag which on the average contains 5 to 7 percent total copper, of which 3 to 4 percent is metallic copper.

In experiments reported by Banks (53), it has been demonstrated that substantial amounts of copper-, nickel, tin, and zinc can be recovered from secondary blast furnace slags. Pilot scale test work has shown that between 50 and 70 percent of the copper contained in a blast furnace slag could be recovered by the addition of coke breeze to the slag while it was maintained in a molten condition in an arc furnace. The copper was recovered as a metal phase assaying between 30 and 40 percent copper and 30 to 60 percent iron.

In other more recent work conducted by the Bureau of Mines (63), bench and s-11-scale tests have shown the technical feasibility of recovering a recyclable copper product from a waste granulated secondary blast furnace slag. Both gravity separation and flotation on a bench scale produced suitable copper concentrates containing more than 30 percent copper and recovering more than 60'percent of the copper. Gravity concentration required grinding the ore through 35 mesh; flotation required grinding the ore through 65 mesh.

In the same report a small-scale continuous test was conducted to investigate simple gravity-separation techniques. A 40-hour test was made with a circuit that included rod mill grinding, a Humphreys sprial, and shaking tables. Two concentrates were prepared according to particie size; the plus 35 mesh concentrate contained 68 percent copper, andtheminus 35 mesh concentrate contained 72.5 percent copper. The composite of the copper products contained 70 percent copper; the process recovered 60 percent of the total copper and 90 percent of the metallic copper. Copper slags contain large amounts of iron oxide which are not easily separated by physical means. On remelting, the slag usually attacks oxide refractories because of the iron. It was found that if the liquidus temperature of a coppper slag is lowered, not only is the energy consumption for melting reduced, but refractory corrosion is also minimized. In one process (57) the melting temperature of the slag was actually lowered from 1,-5OO0C to 1,250"C by adding coal or graphite into the batch.

In this experiment (57) iron was recovered from the slag, and glass fibers were produced from the iron-depleted melt. The mix consisting of 90 percent slag and 10 percent CaC03 was melted with 9 percent of coal powder at 1,350"C for three hours in an electric furnace. Iron metal formed and sank to the bottom. The glass fiber was drawn from the upper portion of the melt.

Operations for recovering copper from reverberatory slag at White Pine Copper Division smelters in Michigan are described by the authors of Reference 58.

Their first step was to crush and screen the slag to 1-inch size and then fed to dual-density heavy media separation systems. A 3.0 specific gravity slurry was maintained in the first compartments. The 3.0 float was too lean in copper for further processing and therefore removed. The 3.0 sink was collected, drained, and put into the second heavy media slurry chamber which was maintained at 3.4 specific gravity. The 3.4 float was stockpiled for future processing while the 3.4 sink constituted the heavy media copper concentrate which was resmelted.

In addition to the copper concentrate produced, the 3.0 float and jig tails from the crushing and screening operations, made an excellent aggregate for a variety of construction purposes including road base, drain fields, railroad ballast, concrete, and asphalt roads. Between the years 1972 and 1977, 514,000 tons of these waste products were used in aggregate applications (58).

Also reported are the sales of sized dump slag being sold to mineral wool manufacturers at rates up to 1,000 tons per week. The slag is mixed by the manufacturer with other materials to adjust the overall composition and melted with coke in a cupola furnace. The molten stream issuing from the . furnace is spun into mineral wool to be used as insulation (58).

4.3.2 Metal Recovery from Dusts

At the Kosaka Mine and copper smelting operations in Japan, all valuable metals in the smelter dusts have been recovered since 1975 (259). The basic steps in the process are leaching of the dusts to recover lead sulfate as leached residue, copper recovery from the solution, iron arsenate precipitation as the most stable form for arsenic disposal, cadmium recovery as sponge cadmium and finally zinc recovery as zinc hydrate from the solution. The primary purpose of the leaching step is to dissolve as much as possible of the soluble metals in the source materials and consequently to raise lead sulfate purity as high as possible. Two stage leaching is adopted since the dust needs strong acid leaching.

Copper recovery takes place from the leached solution. The process consists of three steps: CuS precipitation by H2S, neutralization by CaC03 and cement copper precipitation by iron powder.

The solution after copper removal is further neutralized by CaCO 3 until pH 4.8 is reached. Keeping the condition, iron and arsenic in the solution are oxidized by blowing air to form a stable compound iron arsenate. This has a good property for precipitation and filtration.

Cadmium recovery from the solution is carried out by zinc dust purification method that is commonly used for zinc electrolyte purification. Cadmium is recovered as sponge cadmium. It is also possible to recover cadmium as a form of cadmium sulfide by adding hydrogen sulfide if desired. Zinc is recovered as zinc hydrate by ammonia neutralization.

In another report presented by representatives of the Sumitomo Metal Mining Company of Japan '(67), their Kunitomi copper smelter has been recovering bismuth from its smelting dust since 1968. The bismuth is concentrated in the electrostatic precipitator dust and the converter dust.

Bismuth contained in the converter dust is leached by a mixture of salt and sulfuric acid, and is then recovered by a concentration process using scrap iron. The bismuth cake thus recovered and the electrostatic precipitator dust are reduced to crude bismuth metal which is further refined to a high purity bismuth metal by the gyro-metallurgical refining process.

At the Rokana copper smelter in Zambia, processes have been developed for recovering copper and bismuth from the reverberatory and converter dusts (61). Work was conducted on two processes--sulfuric acid leaching and arc furnace smelting--to extract copper from reverberatory dust. Both processes offer potential treatment routes. The acid leaching process is a simpler approach, but results in lower copper recoveries than arc furnace smelting.

A process that incorporates acid-chloride leaching has been developed for recovery of both the copper and bismuth from converter dust. The bismuth is recovered as bismuth oxychloride, which is smelted to produce a crude metal that contains 98 to 99 percent bismuth. The process is capable of recovering 90 to 95 percent of both copper and bismuth from the converter dust (61).

A pilot plant for treating 385 pounds per day of converter dust was constructed and operated successfully for several months and demonstrated the technical feasibility of the process (61).

107 ..

4.3.3 Miscellaneous Recycling

Successful mini plant and laboratory test work has indicated that high purity selenium can be produced from copper refinery slimes (68). The recovery technique fits into existing unit operations, and the selenium recovery was close to 99 percent.

Inco Metals Company's copper refinery at Copper Cliff, Ontario produces copper anode slimes containing between 10 to 20 percent nickel oxide along with normally occurring selenides and tellurides. Because of the copper-nickel association in the ore body, significant amounts of platinum group metals are also present in the slimes. Anode slimes are mixed with concentrated sulfuric acid and continuously digested in an electrically heated reactor at 210 to 215°C for 30 minutes. The product is water leached to extract about 97 percent of the copper and nickel, 85 percent of the tellurium, and 5 to 10 percent of the selenium. Water leach residue typically consists of: Se (20,to 35 percent); Ag (20 to 30 percent); Pb (10 percent); Si02 (10 percent); Cu+Ni (less than 5 percent); and Te, Au, Pt, Pd, etc. (less than 1 percent).

A simplified description of treating the sulfation residue has the following major steps:

o One stage caustic pressure leaching o Two stages of solution purification

o Selenium reduction * o Preparation of si1 concentrate by dilute sulfuric acid leaching

Commercial grade selenium and a precious metals concentrate suitable for smelting'are the two products of the process (68).

Sulfuric acid-hydrogen Peroxide pickle liquors are being used more and more in the U.S. and Europe as an alternative to dichromate. One of the main advantages of peroxide over dichromate is the ability to regenerate spent liquors and to recover copper electrolytically or by cooling crystallization of copper sulfate. Copper recovery by electrodeposition and also regeneration of the sulfuric acid has been accomplished for many years with simple sulfuric acid pickling solutions. This method is suitable also for peroxide pickles with low peroxide centration (56).

As with the regeneration of simple sulfuric acid pickles, thin starter sheets of copper are used as cathodes and thicker sheets of anti- monial lead as anodes. The cell is operated at a cathodic current density of 20 to 40 A/m2 until the copper content is reduced to about 3g/l. By filtration of the pickle liquor, insoluble contaminants such as lead sulfate, which may have a harmful effect in pickling of brass can be removed (56).

108 4.4 ZINC INDUSTRY

4.4.1 Metal Recovery Processes

Primary smelting of zinc in horizontal retort distillation furnaces leaves a residue consisting of the following: unburned carbon, iron as metal shot and stringers, copper inclusions in the iron as metal and as sulfides, silver in solid solution in the copper, and glassy slag. After removal from the retorts and cooling, the residue is crushed and magnetically separated to produce an upgraded copper-silver material that will economically justify shipment to a copper smelter for the recovery of these two metals. If the magnetic separation fails to achieve necessary enrichment, the residue is stockpiled. When the product.is salable, payment is for copper and silver only, zinc and iron being lost as they are not recovered by the copper smelter (242).

According to the research conducted by the Bureau of Mines (242), recovery of essentially all of the zinc, copper, silver, and iron contained in the magnetic fraction of horizontal zinc retort residues appears technologically feasible by either reduction smelting or by matte smelting. By either method, essentially all of the zinc is recovered as the oxide.

The following conclusions were reached in the above mentioned report .

Smelting under reducing conditions with no addition of sulfur or flux yields a small quantity of rich matte and a high-copper iron. Vir- tually all of the silver, part of.the copper, and a small amount of the iron is contained in the matte; the remainder is in the high-copper iron alloy. The alloy, when chilled rapidly, is amenable to granulation by crushing, or it niay be shotted by being poured through a jet of water. These are materials that can be substituted for scrap cans in the cementation of copper from leach solutions. Copper and silver would be recovered from both matte and the consumed iron.

Matte smelting with pyrite produces a matte containing essentially all of the silver, about 80 percent of the copper, and 37 percent of the iron; a co-product is an iron-copper alloy containing all of the remaining iron and copper.

Extraction with metallic lead recovers essentially all 'of the copper from the molten iron-copper alloy with little or no loss of iron. If copper sulfides are present in the iron-copper metal, calcium carbide can extract the sulfur in a slag and liberate copper for subsequent dissolution in metallic lead. The decoppered, desulfured iron would be an acceptable foundry iron.

Treatment of the molten copper-iron alloy with sodium sulfate reduced copper and sulfur from 6.28 and 2.75 percent to 0.05 and 0.08 percent, respec- tively. For this range of refining, iron loss to the slag was about 25 percent.

109 In a paper describing the development of a small commercial facility located in Laramie, Wyoming, to produce zinc sulfate monohydrate from zinc smelting flue dust, some very positive' results are presented (240, 241).

In the plant, the zinc flue dust was conveyed from the stockpile into an agitated slurry tank. There it was slurried with dilute sulfate solution from the wash filter or barren liquor from solvent extraction. The slurry was pumped to a heated air sparged leach tank where it was agitated with a retention time on the order of eight hours.

Exposure of the slurry to an oxidizing environment converted soluble ferrous iron to insoluble ferric iron which, upon pH adjustment and settling and filtering, reduced the iron content of the zinc sulfate solution to a negligible level. An additional advantage of air sparging was that as the ferric iron was removed, other impurities such as arsenic, were co-precipitated.

Slurry was continuously pumped to thickeners where pregnant liquor was separated from the insolubles. Simple filtering methods were employed such as a pre-coat vacuum filter or sand bed to remove suspended solids. The filter cake was repulped with water and recirculated to either leach or thickners. Thickener underflow (which contained lead, and some silver and cadmium, etc.), was stockpiled for future processing.

The clarified pregnant liquor was then subjected to further purification. The liquor was pumped to an agitated tank where zi dust was mixed with the liquor to replace copper, cadmium, and silver ions This slurry was then filtered through pressure filters to remove the solid impurities. The purified zinc sulfate solution (12 percent contained zinc) can be marketed as solution grade or can be used as feed for the spray dryer and granulator.

The spray dried zinc sulfate monohydrate was a free-flowing, granular white solid. The product had a particle size distribution of 2 percent, 325 mesh, 65 percent 4-200 mesh, which was comparable to other commercial products. The product was marketed by distributors to both agricultural and industrial sectors.

In other research work it has been demonstrated on a laboratory scale, as feasible to recover zinc oxide from slag produced by the Imperial Smelting Process (239). This smelting process is based OR a blast furnace operation in which zinc-bearing ore concentrates are reduced by carbon at high temperature. At one end of the blast furnace a slag is produced. The composition of the slag varies with the type of gangue present in the ore concentrates; it can contain up to 10 percent by weight of zinc oxide.

The method devised to concentrate and recover the zinc oxide present in the slag is presented in Reference 239. It is based on a pre- liminary refiring cycle, which enables the concentration of almost the whole of the zinc available in the slag in a newly formed phase, a magnetite-franklinite solid solution.(Fe304-ZnFe204), embedded in a

110 glassy ground mass. Sulfuric acid leaching treatment was found to exhibit a selective dissolving action, leaving most of the zinc-bearing phase as a solid residue.

An overall recovery of about 85 percent of the total zinc available in the original slag was obtained after 8.5 hours of leaching action of a 2 percent sulfuric acid solution at at temperature of 100°C.

4.5 LEAD INDUSTRY

4.5.1 Me tal Recovery Processes

The Cominco smelter, located at Trail, British Columbia, Canada, produces 200,000 tons of lead bullion annually in conventional blast furnaces. The resulting slag from these furnaces, 231,400 tons annually, containing 18 percent zinc and 2.5 percent lead, is treated directly in the slag fuming plant for the recovery of these metals and other valuable metals present in smaller quantities (155).

Slag fuming is carried out on a 60 ton batch process in completely water-jacketed furnaces where the slag is blown with a mixture of powedered coal and air. The coal-to-air ratio is controlled to maintain strongly reducing conditions, thus, fuming the metals from the slag. The metal va- pors are subsequently re-oxidized with air above the bath. The mixture of oxides is carried from the furnace by the gases through waste heat boilers and cooling flues to a baghouse where they are controlled.

The slag fuming plant, under normal conditions, will treat 770 tons of slag daily. The throughput in the plant is dependent on the fuming rate of zinc, which is a function of the composition of the slag, and on the ecomomics, which dictate the extent to which zinc is fumed from the slag. Zinc and lead collected daily as fume average 125 tons and 19 tons, respec- tively, corresponding to recoveries of 89 percent zinc and 98 percent lead from the slag.

In other research work, it was found that on a pilot plant scale, copper could be recovered from copper-lead dross, a by-product of the refining of lead bullion produced by the Imperial Smelting Process (55).

This dross contains some 27 percent copper, 50 percent lead, 4 percent zinc, and one percent sulfur. An ammonium carbonate leach at atmos- pheric temperature and pressure removed the copper from the dross while the lead zinc and precious metals remained in the leach residue for return to the smelter. The leach solution containing the copper was then purified by liquid-liquid ion exchange using LIX64N in a kerosene dilutent. (LIX is the regestered trademark of a range of hydroxy oximes marketed by General Mills, Inc.) The pilot plant had a capacity of 1.5 pounds per hour of CuSo 4' 5H20 and was operated for some 50 days on a two shift basis. Copper sulfate of excellent purity was produced, and the results obtained indicated that a copper recovery of 90 percent from the dross could be expected from a commercial installation.

111 5.0 SELECTED RESOURCE RECOVERY PROCESS AND WASTE MANAGEMENT APPROACHES

The preceding sections of this report have presented descriptions of the various smelting and refining industries and the past, present, and proposed resource recovery techniques available to the industry. This has included descriptions of each resource recovery process, its stage of development, the waste streams to which it can be applied, and the material that can be recovered. This section will highlight selected resource recovery practices or processes which have potential for greater application i within the industry. This is not to say that these waste utilization techniques will be advanced by increased awareness, but that these selected utilizations are among those which have the least amount of disadvantages to overcome.

Section 5.1 presents those processes identified in Section 2 that are supported by significant research and/or proven experience. It is felt that these could be further utilized within the industry if certain conditions changed. The feasibility of using a specific waste resource depends on a number of factors, many of which are interrelated. These resource recovery processes were evaluated for technological, economic, regulatory, and institutional factors and were in general, found to be favorably influenced by one or more of these factors. However, generally no option will be widely implemented unless either it is economically profitable, or required by law.

Section 5.2 identifies those resource recovery processes that FAL feels could potentially have greater application within the industry but are only supported by 1imited.research or applications. These processes are under active investigation and each could potentially reduce the waste accumulations, or the costs of treatment and disposal.

5.1 KEY AREAS FOR INCREASED RECOVERY

There are three ways to increase the state-of-the-art of resource recovery of the smelting and refining industry:

o Increasing the level of development of recovery processes o Expanding the transferability of waste or recovered materials o Developing additional recoverg/utilization processes or identifying additional waste streams from which materials may be recovered

In order to successfully expand current resource recovery/waste utilization practices, these methods must be applied to those areas with the greatest likelihood of success. These three methods describe how the state-of-the- art can be advanced, but they alone do not determine which resource recovery processes have the best potential for being advanced. Additional factors must be considered in selecting those processes that have the greatest potential. Each of the 59 resource recovery processes or practices reviewed in this study have varying degrees of potential for advancement. However, some of the processes have more potential, as mentioned earlier, in the form of fewer disadvantages to overcome. This section presents those processes selected felt to be representative of resource recovery processes in the smelting and refining industry which exhibit fewer disadvantages, and thus represent the best potential for further use. Those selected are:

0 Coke plant tar used as a fuel 0 BOF slag used as an aggregate 0 Steel making dust recycled to the blast furnace 0 Electric furnace dust recycled to the blast furnace 0 Open hearth furnace dust recycled to the blast furnace 0 Metal recovery from red mud 0 Red mud used in construction 0 Red mud as in SO2 absorbent 0 Metal recovery from copper slag 0 Metal recovery from zinc smelter residue 0 Copper recovery from lead smelter dross

The reason for such a short list of processes with potential for advancement is that waste recycling in the smelting and refining Industry 1s practiced widely. Much of the high volume wastes generated are being utilized within the plants, or enjoy strong market demands on the outside. For example, blast furnace slags are generated at a rate of approximately 30 million tons per year. Yet in.some smelting plants very little slag is ever accumulated, due to its recyclability and use as an aggregate. There- fore, the resource recovery processes selected here for reasons of potential increased use, were selected because they have not been accepted’ or practiced widely, but have been proven to be at least technologically feasible. i Emphasis was placed on those wastes which are normally landfilled. The selection of these resource recovery processes, based on the consideration of four general factors, will be examined in the following pages. These factors used to assess the advantages and disadvantages of I each process are: !I 1 ! o Technological o Economic o Regulatory o Institutional ! I Included In these four general areas are such factors as: application of the process to more than one waste stream; quality and volume of research suggesting the potential for advancement; stage of process or technology development outside the United States; and current and future market for recovered materials or waste-derived products. Thus, the advancement potential for each process can be defined by examining each of these general analysis areas (selectson criteria) within the context of the available information. 8 This section and following subsections present the analysis of these resource recovery and waste utilization areas with regard to the four selection criteria. These criteria are described in more detail below.

5.1.1 Technological Selection Criterion

The technological evaluation of a process primarily determines whether or not the technology is developed, or significantly far enough along, to successfully recover or utilize waste materials. This evaluation includes the use of the process in the United States as well as the successful technological development of the same, or similar process, in other industries or countries. For example, a technology may be commonly practiced in Europe, but not implemented nor researched in the United States. Under these circumstances, the technology would have good potential for advancement through implementation in the United States. Even if the technology exists or has been proven, there may be other technological considerations affecting its potential. These considerations include, but are not limited to, adaptability of present equipment, changes in product quality, limited process applications, or equipment problems due to waste streams being treated. Such factors can either enhance or prevent the use of a technologically proven system. Figure 5-1 presents the various technological advantages and disadvantages that were evaluated for each resource recovery or waste utilization process'in order to assess the overall influence of this. criterion. Studying Figure 5-1, it'can be seen that all of the processes are technically feasible and that the advantages generally outweigh the disadvantages. This corresponds to the need for a process to be technologically feasible in order for it to be considered for implementation within the industry. Few establishments will opt for a technologically unproven or pmblem-ridden process for resource recovery when other proven waste management practices exist.

These processes are the most likely to experience continued or : more widespread utilization in the future, due to the extent of their demonstratable technology. 5.1.2 Economic Selection Criterion

The main purpose of the economic evaluation is to illustrate whether or not the process is economically feasible. This considers cost competitive- ness with conventional materials or with conventional waste. treatment/disposal techniques. Also considered are factors such as additional capital investment required and market and marketing conditions.

114 I I I II I Steel- ke Tar WF Slag makina Dust bcycled to Anal ria Crlterla Fuel A re ate Blast Furnacd

Technology Exists Used in the U.S. Used In Ocher Coun cr f as 0 I 1 I Applicable to I Majority of Plants o 0 I X

Technology Unproven on a Cowrcial Scllt I I

-KEY: o Criterion applies HA Criterion doer not apply x by or MY not be true deprndlng upon proccsa utilized Source: Franklin Assocfatts, Ltd.

Figure 5-1. Technologilcal Advantages and Disadvantages of Selected Economics plays a very decisive role in the undertaking of any project. This is especially so when economic benefits are vague and partially unpredictable which is quite often the case in implementing new technologies. Even though a process may be technologically proven, it will rarely be considered unless it can provide a savings of some sort. The savings can be in the form of marketability of the recovered material, reduced disposal costs, reduction in raw material requirements or some combination of these three factors.

As can be seen in Figure 5-2, economic feasibility must, in most cases, be determined on an individual basis. A recovery process that operates economically at one plant may not work at another. Small variations in smelting or refining processes or equipment can make some recovery technologies economically unfeasible.

Also, economics are relative to individual plant locations and situations. Smelting and refining plants located in areas where land values are not very high will feel no urgent economic reasons for recycling or utilizing their waste products when they can be stored or im- pounded on relatively inexpensive land.

Thus, the economic advantages of the processes presented in' Figure 5-2 are not as clearly defined as are the technological advantages. However, several of the selected resource recovery processes do enjoy economic advantages in particular instances and, coupled with their technological advantages, are currently being practiced in the industry. These processes are:

o Coke tar as fuel o BOF slag as aggregate o ' Steelmaking dust recycled to the blast furnace o Electric furnace dust recycled to the blast furnace o Open hearth furnace dust recycled to the blast furnace o Metal recovery from copper slag o Red mud SO2 absorbant

These resource recovery processes listed above are being practic on a limited basis. As mentioned previously, economics is the crucial factor in most cases and thus limits the widespread utilization of these recovery processes. The reason they are included here is because of their future potential to spread if and when certain conditions change relevant to plant requirements for implementation.

116 ... ..

X Processes I I Y I I Uneconomical for I Small Facilities I I Io I

-KEY: o Crlterion applies ? Unable to deternine fr~pdata x May or may nut be true depending upon individual practicer and processes y Xay or may not be true depending upon distance from waste source Source: Frankllo Assaclater. Ltd.

Figure 5-2. Economic Advantages and Disadvantages of Selected Resource Recovery Processes ..

5.1.3 Regulatory Selection Criterion

Federal, state, and local regulations can limit the use of a resource recovery process or a recovered material. These regulations may specify contaminent levels in products or the quality of the material used in other products. Also, installation of pollution control devices for resource recovery processes may be required in order to meet emission, health, and safety standards.

Regulations can also indirectly foster the use of a resource recovery process. The regulations may make current disposal practices uneconomical, thus forcing the waste generator to either change the disposal practice or recover the materials in the waste. Also, regulations may make resource recovery more cost-effective than current disposal practices. In addition, reduction of solid and liquid waste streams through resource recovery may significantly reduce the impact of applicable regulations.

Figure 5-3 summarizes the advantages and disadvantages associated with the individual resource recovery processes, due to various regulations. A key issue is whether or not the process or product is of regulatory concern at either the Federal, state, or local levels. The absence of regulatory concern can be viewed as a clear advantage to a resource process since this eliminates the need to conform to regulatory requirements.

In the metal smelting and refining industry, many of the resource recovery processes utilizing waste streams recycle the wastes back into the production process and are thus not independently subject to regulatory control. Product grade and specifications play the major role in determining how much waste can actually be recycled back into production.

However, in cases where the waste product is removed from the plant and utilized elsewhere, state and local regulations may apply. For example, using BOF slag as aggregate, or red mud in construction products, will be subject to state and local building codes which contain mix, strength, and structural specifications.

5.1.4 Institutional Selection Criteria

Industry attitudes toward resource recovery practices depend on a number of factors including: favorable or unfavorable impression of success created by the other three selection criteria, resistance to change, and the specific nature of che smelting and ref.ining industry. in generai, the only way a change in current waste management practices will occur is if the other three criteria (technological, economic, and regulatory) are positive. These factors largelygovernan industry's perception as to whether or not a recovery process will be successful. This perception, in turn, influences the willingness of an industry to try a recovery process. For example, if a resource recovery process is percesved as not being successful based on potential or actual technological or economic problems, then industry is likely to be unwilling to implement the process whether or not the perception itself is accurate.

118 a

Product or Process of Regulatory Concern X 0 0 \ @ Requlres Pollution t- t- ControlIEqulpment 0 0 0 , -KEY: o Criterion applies ? to Ifrom Unable determine data x by or may not be true depending on process utilized

Source: Franklin Associate., Ltd. .-

Figure 5-3. Regulatory Advantages and Disadvantages of Selected Resource Recovery Processes

I. ,

-KEY: o Cr,iiLerionl applies ? Unable to determine from data x Hay or may not be true depending upon individul plant

Source: Franklin Associates. Ltd. i

Figure 5-4. Institutional Advantages and Disadvantages of Selected Resource Recovery Processes Attitudes of industry personnel and local residents also affect the implementation of resource recovery processes. At the management level, there may be resistance to changing or replacing proven processes or materials with newer ones, especially if the older processes and materials have been used for a long time. Such situations may be perceived by the workers as a threat to their job security. Also, implementation of new processes may require workers to learn new jobs or processing techniques. Additionally, residents near an industrial facility may object to the installation of new equipment, especially if such equipment is perceived as a source of substances that may threaten their health or the surrounding environment.

Additionally, there is a natural reluctance on the part of industry engineers to use materials that by their very name, "waste materials,'' imply that they are inferior to conventional materials. On the other hand, one must guard against the enthusiastic and sometimes exaggerated claims of researchers who have investigated the process in laboratory studies.

Attitudes toward resource recovery and waste utilization depend on the overall influence of the other three criteria (technological, economic, regulatory) as well as a number of subjective considerations. Figure 5-4 presents the institutional factors that were evaluated to determine the overall influence of' this criterion on the advancement potential of the selected resource recovery processes.

From Figure 5-4 it can be seen that there has been a limited willingness to incorporate resource recovery processes. This is a reflection of the fact that some companies are more progressive than others in their concerns about waste recycling. A recovery process has a much better chance of succeeding if the plant operators are determined to make it work. This determination to succeed factor was an underlying key in each of the successful applications of waste utilization processes reviewed in the literature. Thus this institutional criteria can be considered as important as the preceeding three criteria; technology, economic, and regulatory.

5.1.5 Summary of Analysis

This section presents a summary of the analyses of the key resource recovery and waste utilization areas with regard to the four selection criteria. The results of this analysis are presented in Figure 5-5. This figure illustrates the influence of the selection criteria on the potential for greater application and development of these selected processes.

This influence may be positive, negative, or even variable depending on specific industry or process parameters and characteristics. The overall potential of a recovery' process depends on all four factors, thus a

negative influence of one factor does not necessarily preclude the process from having good advancement potential. _Although these four selection criteria have been largely discussed independently, they are in fact interdependent. The institutional selection criteria is largely affected by the other three criteria. The other three factors (technological, economic, and regulatory) are also interdependent. For example, regulatory controls on emissions can result in technological problems in trying to adapt the system to meet these regulatory standards. This in turn can increase the cost of the system to the point that it is no longer economical.

Analysis of Figure 5-5 reveals that some of the selected resource recovery processes have an advantage for continued or increased . utilization.' Furthermore, it can be seen that with a change for the positive in just one of the four evaluation factors, many of the selected processes could realize greater utilization potential. As noted earlier, the economics of a particular process plays the heaviest role in implementation decisions, and the economics of most of the smelting resource.recoyeu processes.must be evaluated on an individual basis.

Another point to notice in Figure 5-5, is that none of the recovery processes scored a plus in every criteria, which would indicate widespread use of the waste, resulting in very small accumulations of the waste product. These recovery processes were chosen because the waste products they utilize are not being completely consumed, and thus stockpiles are accumula- ting. Therefore, these recovery processes represent future potential for the recovery of as yet, under utilized waste materials.

5.2 EtERGING TECHNOLOGIES WITH GOOD POTENTW

In addition to the selected recovery processes presented in Section 5.1, there are several others that are felt may prove valuable to the industry in the future. These processes are:

o Steelmaking slag as coal spoil acid neutralization o Steelmaking slag as an SO2 absorbant o Steelmaking fines as industrial pigments

The first has reached the stage of a small scale field trial, the second is still under review in research studies, and the third'has reached full scale utilization in Australia. These three have the potential for widespread use resulting in raw material savings, pro-vidiag the initial favorable results can be duplicated on a commercial scale in the United States. Each is discussed in the following pages.

5.2.1 Steelmaking Slag as Coal Spoil Acid Neutralization (197)

Experiments conducted at the Butler County Community College, Butler County Pennsylvania (197), have shown on a laboratory scale, that treating

123

I coal mining and washing wastes with steelmaking slag, can neutralize the acidic properties of these wastes. The results of water leaching various combinations of spoil-slag mixtures showed that all of the drainages were neutralized after a few weeks of reaction and all were relatively free of heavy metal leachates.

The indication is that steelmaking slag, layered or mixed into the spoil itself, can surpassiand negate the formation and release of acid as it occurs. One cubic yard of slag can apparently cope with the acid potential of four to six cubic yards of spoil. Although the sources of steelmaking slag tend to be remote from the sites of coal production, they are connected, at least theoretically, by a steady stream of empty coal- carrying vehicles.

If further planned tests of this process succeed, there is the potential for reducing a very,considerable current expense for acid treatment operations.

5.2.2 Steelmaking Slag as an SO:, Absorbant (197)

In a Brookhaven National Laboratory report, BNL 50891, entitled "Regenerative Process for Desulfurization of High-Temperature Combustion and Fuel Gases-Quarterly Progress Report No. 9," it was stated, in part, that several calcium silicates had higher sulfation rates and sulfation capacities than pure burnt lime, a common substance used for sulfur ab- sorbtion. In addition, the silicates were easier to regenerate than lime. The silicates used in the experiments were costly, synthesized reagent chemicals. Since then the Brookhaven people have looked at steelmaking slags as an abundant and cheap supply of calcium silicate.

It,is a use still to be proven, but Brookhaven has undertaken to study the use of steelmaking ,slags for stack sulfur control. If a practical technology can be devised, it would represent a world wide waste utilization capability.

5.2.3 Steelmaking Fines as Industrial Pigments

A full scale utilization of steelmaking fines for pigments in the cement and paint industries, by an Australian firm, is described in Re- ference 175.

Accnrding to the Australian company which reported the facts, sales and acceptance of these alternative pigments have been successful.

The brown pigment is derived from steelmaking fume produced in the basic oxygen steelmaking process. The red is derived from iron oxide produced in the regeneration of hydrochloric acid liquor used for steel pickling. The black can be derived from any fine iron oxide dust by reduction roasting to form magnetite which is then processed to a fine black pigment. The yellow can be derived from neutralization of sulfuric acid liquor which is used for pickling steel in some plants. These pigments are new grades based on synthetic iron oxides, and strict quality control procedures have been developed to ensure con- sistent color grading.

If this technology could be imported to the United States, it would represent an additional opportunity for the utilization of iron oxide wastes which are not now being completely recycled.