DETAILED ACTION
Notice of Pre-AIA or AIA Status
The present application, filed on or after March 16, 2013, is being examined under the first inventor to file provisions of the AIA .
Priority
Receipt is acknowledged of certified copies of papers required by 37 CFR 1.55.
Information Disclosure Statement
The information disclosure statements (IDS) submitted on 10/22/24 and 8/15/24 were filed. The submission is in compliance with the provisions of 37 CFR 1.97. Accordingly, the information disclosure statements have been considered by the examiner.
Drawings
The drawings were received on 1/24/24. These drawings are acceptable.
Claim Rejections - 35 USC § 103
In the event the determination of the status of the application as subject to AIA 35 U.S.C. 102 and 103 (or as subject to pre-AIA 35 U.S.C. 102 and 103) is incorrect, any correction of the statutory basis (i.e., changing from AIA to pre-AIA ) for the rejection will not be considered a new ground of rejection if the prior art relied upon, and the rationale supporting the rejection, would be the same under either status.
The following is a quotation of 35 U.S.C. 103 which forms the basis for all obviousness rejections set forth in this Office action:
A patent for a claimed invention may not be obtained, notwithstanding that the claimed invention is not identically disclosed as set forth in section 102, if the differences between the claimed invention and the prior art are such that the claimed invention as a whole would have been obvious before the effective filing date of the claimed invention to a person having ordinary skill in the art to which the claimed invention pertains. Patentability shall not be negated by the manner in which the invention was made.
The factual inquiries for establishing a background for determining obviousness under 35 U.S.C. 103 are summarized as follows:
1. Determining the scope and contents of the prior art.
2. Ascertaining the differences between the prior art and the claims at issue.
3. Resolving the level of ordinary skill in the pertinent art.
4. Considering objective evidence present in the application indicating obviousness or nonobviousness.
This application currently names joint inventors. In considering patentability of the claims the examiner presumes that the subject matter of the various claims was commonly owned as of the effective filing date of the claimed invention(s) absent any evidence to the contrary. Applicant is advised of the obligation under 37 CFR 1.56 to point out the inventor and effective filing dates of each claim that was not commonly owned as of the effective filing date of the later invention in order for the examiner to consider the applicability of 35 U.S.C. 102(b)(2)(C) for any potential 35 U.S.C. 102(a)(2) prior art against the later invention.
Claims 1-10 are rejected under 35 U.S.C. 103 as being unpatentable over CN 114015885 A (CN'885) in view of CN 117004816 A (CN'816) and CN 111733328 B (CN'328).
As to Claim 1:
CN'885 discloses a method for separating and recycling a waste material containing lithium iron phosphate. (CN’885 Pg. 1).
CN'885 discloses step (1) of subjecting a waste lithium iron phosphate battery to discharging and disassembly to obtain a lithium iron phosphate positive plate. Specifically, CN'885 teaches pre-treatment steps including discharging, crushing, sorting, and sieving waste power batteries to isolate the positive electrode powder matrix. (CN’885 Pgs. 4–6).
CN'885 discloses step (2) of subjecting the lithium iron phosphate positive plate to breaking, followed by high temperature treatment. Specifically, CN'885 teaches breaking and crushing the battery material down to powder fractions. (CN’885 Pgs. 4–6).
CN'885 discloses step (3) of mixing a product with a carbon material, transferring a mixture to a sintering device, introducing a chlorine-containing reagent, and heating and roasting the mixture in a chloridizing atmosphere. Specifically, CN'885 teaches mixing the active substance with carbon powder (such as coal powder) and roasting it with a chlorine-containing reagent at high temperatures (500–1000 °C) to initiate gas-phase volatilization reactions. (CN’885 Pgs. 2–5).
CN'885 discloses step (4) of enabling a gas phase product produced in the roasting process to sequentially pass through a condensation pipeline divided into two steps or stages, where the condensing temperature of the first step is controlled at 200–600 °C to obtain ferrous chloride (FeCl₃) and the condensing temperature of the second step is controlled at 50–170 °C to obtain aluminum chloride (AlCl₃). (CN’885 Pgs. 2–6).
CN'885 discloses step (5) of subjecting a solid phase product obtained by the roasting to water leaching, stirring, and filtration to obtain a lithium-rich solution. Specifically, CN'885 teaches leaching the remaining roasting slag with water or a water solution to dissolve and separate the lithium component from unreacted copper and carbon residues. (CN’885 Pgs. 2–6).
However, CN'885 does not explicitly disclose performing the step (2) high temperature treatment specifically in an air atmosphere; uniformly mixing the materials specifically by ball milling, vacuumizing the sintering device prior to gas introduction, or introducing high-purity Cl₂ gas as the chlorine source in step (3); utilizing preheated, backward-flowing silicon tetrachloride (SiCl₄) as the fluid cooling medium inside the heat exchange tubes of quenching and condensation separation devices 1 and 2 to recover the condensed FeCl₃ and AlCl₃ in step (4); or obtaining a LiCl aqueous solution during leaching and adding sodium carbonate into the solution to recover and precipitate Li₂CO₃ in step (5).
CN'816 and CN'328 disclose the missing limitations of the primary reference. CN'816 discloses gas-phase chloridizing roasting to selectively volatilize and separate iron, explicitly teaching the introduction of a gaseous chlorinating agent comprising elemental chlorine (Cl₂) gas into a high-temperature roasting furnace. (CN’816 Pgs. 1, 3–5).
For the subsequent step-by-step sectional condensation of the gas phase, CN'816 teaches a fractional system configured to pass the gas phase through distinct condensation zones to separate and condense solid iron chloride while dynamically handling and isolating silicon tetrachloride (SiCl₄) fluid streams. (CN’816 Pgs. 1, 3, 5–7).
CN'328 discloses a method for recycling valuable metals from waste lithium-ion batteries by uniformly mixing the active electrode powder material with a carbon material specifically via ball milling to obtain a blended mixed powder prior to feeding the reaction furnace. (CN’328 Pgs. 1–5).
CN'328 further discloses step (5) chemistry where the chloridized roasted product is soaked and leached with water to yield a lithium-rich filtrate, and subsequently adding sodium carbonate into the lithium-rich solution to recover and precipitate a lithium carbonate (Li₂CO₃) product. (CN’328 Pgs. 1, 3–6).
CN'885, CN'816, and CN'328 are analogous arts because they are each directed to the fields of extractive metallurgy and chemical resource recycling. Specifically, CN'885 and CN'328 are directed to the identical field of recycling valuable metals from decommissioned lithium-ion battery electrodes, while CN'816 and CN'885 both utilize gas-phase chloridizing roasting and multi-stage temperature-controlled fractional condensation loops to isolate and purify iron chloride products from secondary material aggregate streams. (CN’885 Pgs. 1–4); (CN’816 Pgs. 1–4); (CN’328 Pgs. 1–4).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to combine the high-temperature chloridizing roasting and fractional condensation process of CN'885 with the ball milling and downstream sodium carbonate precipitation chemistry of CN'328, and further incorporate the elemental Cl₂ gas source and SiCl₄-associated fractional condensation mechanics of CN'816. (CN’885 Pgs. 2–4); (CN’816 Pgs. 3–5); (CN’328 Pgs. 2–4).
A person of ordinary skill in the art would have been motivated to combine these teachings to address the specific drawback identified in CN'885, wherein traditional wet battery recovery causes iron and aluminum to co-leach synchronously with lithium, resulting in a low-purity final output and intensive extraction chemical burdens. (CN’885 Pgs. 2–4).
Utilizing the ball milling step of CN'328 ensures a highly uniform solid mixture, and vacuumizing the furnace prior to introducing the Cl₂ gas of CN'816 is a routine engineering optimization to prevent atmospheric oxygen interference and ensure complete selective chlorination of the metal phases. (CN’328 Pgs. 2–5); (CN’816 Pgs. 3–5).
Furthermore, substituting and managing a counter-current, backward-flowing stream of preheated SiCl₄ through the heat exchange tubes within the broader condensation temperature windows of CN'885 represents a predictable application of the thermodynamic separation data explicitly detailed in CN'816 to cleanly isolate solid FeCl₃ and AlCl₃ fractions. (CN’885 Pgs. 2–6); (CN’816 Pgs. 3–7).
Finally, because pure chloridizing roasting generates water-soluble lithium chloride (LiCl) in the solid slag, concluding the extraction loop by water leaching the product and adding sodium carbonate to precipitate out high-purity battery-grade Li₂CO₃ is a standard and predictable application of the hydrometallurgical recovery steps demonstrated in CN'328. (CN’328 Pgs. 1, 3–6).
As to Claim 2:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (2), the lithium iron phosphate positive plate undergoes a breaking process; and CN'885 discloses that the raw battery material is broken down via an over-crushing and screening system to significantly minimize grain diameter dimensions. Specifically, CN'885 teaches over-crushing the lithium iron phosphate battery material to produce a very fine active powder fraction with particle dimensions restricted to less than 0.1 mm to maximize surface area contact during subsequent roasting. (CN’885 Pgs. 4–6).
However, CN'885 does not explicitly disclose that the lithium iron phosphate positive plate has an average particle size specified as “equal to or less than 3 mm” after the breaking step, since it chooses to define an ultra-fine particle size threshold restricted to less than 0.1 mm.
CN'328 discloses the missing particle sizing limitation of the primary reference. CN'328 discloses a method for recycling valuable metals from waste lithium batteries where the raw positive active material is broken down and processed into a raw active powder. CN'328 teaches that raw battery components are crushed and milled into fine particulate parameters prior to thermal feeding. (CN’328 Pgs. 5–7).
The target parameter constraint of having a broken particle size of “equal to or less than 3 mm” represents a broad, standard operating threshold for crude battery aggregate powders, which is entirely encompassed and inherently met by the far more stringent, fine powder dimensions (less than 0.1 mm) explicitly utilized by the primary reference CN'885 to ensure rapid gas-phase volatilization kinetics. (CN’885 Pgs. 4–6); (CN’328 Pgs. 5–7).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to establish the particle size parameters of Claim 2 within the combined method. A person of ordinary skill in the art would have been motivated to adjust and optimize the mechanical breaking constraints to achieve an average particle size equal to or less than 3 mm because selecting an optimal particle size distribution is a routine variable optimized in metallurgy to facilitate complete gas-solid exposure. Given that the primary reference CN'885 explicitly utilizes an ultra-fine powder fraction of less than 0.1 mm to drive its volatilization reactions, operating at or below a coarser 3 mm standard boundary constitutes a predictable application of the mechanical milling and breaking techniques taught across both battery processing streams to optimize material flowability and reaction rates. (CN’885 Pgs. 4–6); (CN’328 Pgs. 5–7).
As to Claim 3:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (2), the lithium iron phosphate positive plate undergoes a breaking process followed by a thermal operation to facilitate gas-solid chlorination roasting reactions in a tube furnace; and CN'885 discloses that the broken material is heated inside a furnace chamber up to a target reaction temperature of 850 °C to drive the gas-phase separation of the iron and aluminum components. (CN’885 Pgs. 4–5).
However, CN'885 does not explicitly disclose that the step (2) thermal operation is a high-temperature treatment performed specifically in an air atmosphere, at a heating rate of 5–10 °C/min, at a temperature of 400–600 °C, or that the temperature is maintained for 3–6 hours prior to the chloridizing roasting stage.
CN'328 discloses the missing thermal parameters and operational conditions of the primary reference. CN'328 discloses a method for recycling valuable metals from waste lithium-ion batteries where the processed positive electrode material is subjected to segmented thermal processing inside a reaction furnace. CN'328 explicitly teaches implementing precise temperature holds and ramping profiles to regulate the phase transformations of active material mixtures, specifically disclosing that a first-section holding temperature is sustained across a range of 200–500 °C for an explicit duration of up to 3.0 hours to optimize reaction kinetics. (CN’328 Pgs. 2–3, 7–8).
Adjusting heating rates to a standard profile of 5–10 °C/min and managing a thermal pre-treatment window at 400–600 °C sustained for 3–6 hours represents a predictable optimization of temperature ramping and soaking holds taught within the secondary art to guarantee a highly uniform phase configuration before exposing the battery matrix to aggressive gas-phase chlorination. (CN’885 Pgs. 2–5); (CN’328 Pgs. 2–4, 7–8).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to incorporate the specific high-temperature treatment constraints of Claim 3 into the combined recovery method. A person of ordinary skill in the art would have been motivated to configure the step (2) thermal pre-treatment to run at a heating rate of 5–10 °C/min and maintain a temperature of 400–600 °C for 3–6 hours because controlling thermal ramping profiles and isothermal soaking times is a routine variable adjusted in extractive metallurgy to secure complete decomposition of structural components and achieve an optimal phase matrix. (CN’885 Pgs. 2–5); (CN’328 Pgs. 2–4, 7–8).
Combining the baseline chlorination sequence of CN'885 with the step-by-step thermal maintenance profile of CN'328 ensures that the active material reaches thermodynamic readiness in a predictable manner, maximizing the volatility and subsequent extraction efficiency of the valuable target metals. (CN’885 Pgs. 2–5); (CN’328 Pgs. 2–4, 7–8).
As to Claim 4:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (3), the product obtained by the high temperature treatment is uniformly mixed with a carbon material; and CN'885 discloses that the lithium iron phosphate waste material is mixed with carbon (such as coal powder) to establish a reducing atmosphere during high-temperature roasting. (CN’885 Pgs. 3, 5–6).
However, CN'885 does not explicitly disclose that the carbon material comprises at least one specific carbon source chosen from graphite, carbon black, carbon fibers, carbon nanotubes, and amorphous carbon, as it defines the carbon material generally as carbon powder or coal powder. (CN’885 Pgs. 3, 5–6).
CN'328 discloses the missing specific carbon material selections of the primary reference. CN'328 discloses a method for recycling valuable metals from waste lithium-ion batteries where the raw positive electrode material powder is uniformly blended with selected carbon materials. CN'328 explicitly teaches that the battery positive material powder must be ball milled and mixed with specific carbon source materials, explicitly disclosing the use of graphite or carbon black to establish the uniform mixed powder needed for high-temperature reduction roasting. (CN’328 Pgs. 3–5).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to select the specific carbon material parameters of Claim 4 within the combined method. A person of ordinary skill in the art would have been motivated to utilize graphite or carbon black as the carbon material because choosing a standard, highly effective carbon source is a routine variable optimized in extractive metallurgy to ensure strong reduction performance during roasting. Given that the primary reference CN'885 requires a carbon material to establish its reducing atmosphere, substituting or selecting the specific graphite or carbon black carbon sources explicitly taught in the battery-recycling stream of CN'328 represents a predictable application of known raw materials to achieve optimal carbo-chlorination reduction efficiency. (CN’885 Pgs. 3, 5–6); (CN’328 Pgs. 3–5).
As to Claim 5:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (3), the product obtained by the high temperature treatment is uniformly mixed with a carbon material; and CN'885 discloses adding the waste lithium iron phosphate material and carbon uniformly into a mixing ratio configuration. Specifically, CN'885 teaches blending 20 g of the lithium iron phosphate waste material with 1 g of carbon powder (such as coal powder) to generate a uniform mixture prior to high-temperature roasting. (CN’885 Pgs. 5–6).
However, CN'885 does not explicitly disclose that the mass ratio of the product obtained by the high temperature treatment to the carbon material is restricted specifically to a range of 10:1 to 5:1, as the absolute batch amounts detailed in the primary reference establish a calculated mass ratio of 20:1. (CN’885 Pgs. 5–6).
CN'328 discloses the missing mass ratio optimization parameters of the primary reference. CN'328 discloses a method for recycling valuable metals from waste lithium batteries where the raw positive active material powder is ball milled and mixed with a carbon reducing agent. CN'328 explicitly teaches that the carbon material added into the mechanical mix can be optimized across explicit structural mass thresholds, specifically disclosing that the mass of the carbon material is controlled to be 10 % to 30 % of the total mass of the positive electrode active material powder. A carbon material mass range of 10 % to 30 % corresponds directly to raw material-to-carbon mass ratios ranging from 10:1 down to 3.3:1. The claimed range of “10:1 to 5:1” represents a narrow, predictable optimization entirely encompassed within the broad concentration thresholds explicitly taught in the secondary art to guarantee a highly efficient carbothermal reduction. (CN’328 Pgs. 3, 7–8).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to configure the mass ratio constraints of Claim 5 within the combined recovery method. A person of ordinary skill in the art would have been motivated to adjust the mass ratio of the product to the carbon material to be 10:1 to 5:1 because selecting and optimizing the relative proportions of core reactants is a routine variable adjusted in extractive metallurgy to maximize reduction kinetics without wasting excess materials. Given that the primary reference CN'885 utilizes carbon to drive a carbo-chlorination reduction environment, adjusting the mass concentrations to conform to the optimized 10 % to 30 % operating thresholds explicitly demonstrated in CN'328 represents a predictable application of metallurgical raw material limits to ensure a complete gas-phase volatilization of the target metals. (CN’885 Pgs. 3, 5–6); (CN’328 Pgs. 3, 7–8).
As to Claim 6:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (3), a chlorinating agent is introduced into a furnace chamber; and CN'885 discloses that the chlorinating gas stream is passed through the reaction system at a defined, continuous velocity to selectively volatilize the target metal elements. Specifically, CN'885 teaches introducing hydrogen chloride gas into a tubular furnace at an explicit baseline flow rate of 100 mL/min to 120 mL/min for a treated batch sample of 20 g, noting that the introduction speed converts to a normalized baseline of not less than 4 mL/min per gram of raw lithium iron phosphate waste material. (CN’885 Pgs. 3, 5–6).
However, CN'885 does not explicitly disclose that the chlorinating gas is high-purity Cl₂, nor does it explicitly disclose introducing it at a specific volumetric flow rate restricted to a range of 10–50 mL/min, as it defines the chlorinating agent as hydrogen chloride or other alternative compounds moving at an absolute batch velocity of 100 to 120 mL/min. (CN’885 Pgs. 3, 5–7).
CN'816 discloses the missing gas velocity and halogen purity parameters of the primary reference. CN'816 discloses a method for separating components by chloridizing roasting inside a roasting furnace. CN'816 explicitly teaches introducing a gaseous chlorinating agent comprising elemental chlorine (Cl₂) gas to directionalize metal volatilization and isolate pure solid metal chloride fractions. (CN’816 Pgs. 1, 3–8).
Operating a gas delivery line within a volumetric flow rate range of “10–50 mL/min” represents a routine, predictable scale adjustment to accommodate modified sample sizes, smaller reactor design dimensions, or specific gas concentrations while remaining completely within the normalized proportional gas-to-feed velocity bounds, not less than 4 mL/min per gram of raw feed material, explicitly mapped out by the primary reference CN'885. CN'885 (Pg. 3).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to configure the gas delivery parameters of Claim 6 within the combined method. A person of ordinary skill in the art would have been motivated to introduce the high-purity Cl₂ gas at a flow rate of 10–50 mL/min because selecting and optimizing gas delivery rates relative to a given raw material mass is a routine variable adjusted in chemical engineering to maintain steady-state gas-solid interactions and prevent unreacted gas bypass. Given that the baseline chloridizing process of CN'885 establishes a normalized flow-to-mass threshold, scaling the absolute gas input down to a range of 10–50 mL/min when utilizing the pure gaseous chlorine taught by CN'816 represents a predictable application of standard fluid flow limits to optimize selective metal volatilization efficiency. (CN’885 Pgs. 3, 5–6); (CN’816 Pgs. 3–8).
As to Claim 7:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (3), the roasting is performed; and CN'885 discloses that the high-temperature chloridizing roasting operation can be successfully carried out in a segmented, step-by-step, or multi-stage manner. Specifically, CN'885 teaches that the vapor-phase separation process utilizes multi-stage thermal and condensation management zones—expressly defining a first-stage condensation step and a second-stage condensation step—to sequentially isolate distinct metal chloride vapor products from the chlorinated gas stream. (CN’885 Pgs. 2–6).
However, CN'885 does not explicitly disclose that the roasting is performed in exactly two stages with the specific operating thresholds where the roasting in the first stage is performed at a heating rate of 3–5 °C/min and a temperature of 180–300 °C and the temperature is maintained for 1–3 h, and the roasting in the second stage is performed at a heating rate of 5–10 °C/min and a temperature of 350–550 °C, and the temperature is maintained for 3–6 h. (CN’885 Pgs. 2–7).
CN'328 discloses the missing multi-stage thermal parameters and operational holds of the primary reference. CN'328 discloses a method for recycling valuable metals from waste lithium-ion batteries where the blended active material mixture undergoes segmented thermal processing. CN'328 explicitly teaches that the thermal reaction is carried out in discrete sections to optimize extraction kinetics, disclosing that the segmented reduction roasting is a two-stage roasting process: the roasting temperature of the first section is controlled at 200–500 °C for a duration of 0.5 to 3.0 h, and the roasting temperature of the second section is controlled at 600–1000 °C for a duration of 0.5 to 5.0 h. (CN’328 Pgs. 2–3, 7–8).
Setting the first-stage temperature hold specifically to 180–300 °C for 1–3 h and the second-stage temperature hold to 350–550 °C for 3–6 h, while establishing conventional heating ramp speeds of 3–5 °C/min and 5–10 °C/min to transition between these thermal settings, represents a routine, predictable engineering optimization of overlapping thermal processing fields explicitly mapped out in the secondary art to maximize reaction rates. (CN’328 Pgs. 2–3, 7–8); (CN’816 Pgs. 3, 7–8).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to configure the specific multi-stage roasting parameters of Claim 7 within the combined recovery method. A person of ordinary skill in the art would have been motivated to execute the carbo-chlorination roasting in two distinct stages with targeted heating rates and temperature holds because managing multi-step thermal holds and specific ramping speeds is a routine variable adjusted in chemical metallurgy to control the kinetics of step-by-step reduction and volatilization reactions. (CN’885 Pgs. 2–6); (CN’328 Pgs. 2–3, 7–8); (CN’816 Pgs. 3, 7–8).
Integrating the baseline chloridizing process of CN'885 with the precise two-stage segmented furnace framework explicitly demonstrated in the battery-recycling stream of CN'328 represents a predictable application of known thermal schedules to ensure high-purity phase transitions and optimize the selective gas-phase volatilization of the valuable target metals. (CN’885 Pgs. 2–6); (CN’328 Pgs. 2–3, 7–8).
As to Claim 8:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (4), the gas phase product is introduced into a condensation separation device; and CN'885 discloses that the second condensation step or stage is tightly managed within an explicit operational temperature window. Specifically, CN'885 teaches that the gas phase product is passed sequentially into a second condensation zone where the temperature is controlled at 50–170 °C to capture and separate out the aluminum chloride (AlCl₃) product fraction. (CN’885 Pgs. 2–6).
However, CN'885 does not explicitly disclose that before being introduced into the heat exchange tube of the quenching and condensation separation device 2, the SiCl₄ is preheated to 100–120 °C first, or that the SiCl₄ is introduced into the heat exchange tube of the quenching and condensation separation device 2 at a flow rate of 5–10 mL/min, as it describes running the fractional condensation zones generally by structural piping temperature controls without utilizing an injected SiCl₄ preheated fluid stream. (CN’885 Pgs. 2–7).
CN'816 discloses the missing silicon tetrachloride handling parameters and fluid introduction mechanics of the primary reference. CN'816 discloses a method for chloridizing roasting and metallurgical separation where a multi-stage fractional condensation system is used to handle gas-phase chloride mixtures. CN'816 explicitly teaches that silicon tetrachloride (SiCl₄) co-exists dynamically as a fluid stream alongside volatilized metal chlorides within these targeted gas-phase pathways. (CN’816 Pgs. 1, 3, 5–7).
CN'816 further teaches that the multi-stage condensation loop includes a distinct step where the roasting gas phase is selectively condensed across a second-stage temperature profile controlled at 25–30 °C to manage phase distribution. (CN’816 Pgs. 3, 5–8).
Preheating the SiCl₄ medium to a specific boundary of 100–120 °C represents a direct, predictable configuration to align the heat-exchange fluid with the broader 50–170 °C aluminum chloride capture window explicitly taught by the primary reference CN'885 to maintain strict local thermodynamic equilibrium, while running the fluid introduction line at a routine volumetric flow rate of 5–10 mL/min represents a standard scale adjustment to maintain stable counter-current cooling velocities inside the heat exchange tubes. (CN’885 Pgs. 2–6); (CN’816 Pgs. 3, 5–8).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to preheat the SiCl₄ fluid to 100–120 °C and introduce it into the heat exchange tubes of device 2 at a flow rate of 5–10 mL/min within the combined method. A person of ordinary skill in the art would have been motivated to preheat and regulate the delivery rate of the SiCl₄ stream because managing the temperature and mass flow of an injected cooling fluid is a routine engineering choice adjusted in gas-phase fractionation to maintain sharp condensation boundaries and prevent gas pipe fouling. (CN’885 Pgs. 2–6); (CN’816 Pgs. 3, 5–8).
Combining the dual-condensation layout of CN'885 with the SiCl₄ gas metallurgy and condensation properties explicitly demonstrated in CN'816 represents a predictable application of known fluid mediums to tightly control the 50–170 °C fractional thermal zone, enabling the clean, steady-state capture of solid aluminum chloride fractions away from the downstream lithium residues. (CN’885 Pgs. 2–6); (CN’816 Pgs. 1, 3, 5–8).
As to Claim 9:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (4), the gas phase product is passed through a condensation separation device; and CN'885 discloses that the first condensation step or stage is tightly managed within an explicit operational temperature window. Specifically, CN'885 teaches that the gas phase product is passed sequentially into a first condensation zone where the temperature is controlled at 200–600 °C to capture and separate out the ferrous chloride (FeCl₃) product fraction. (CN’885 Pgs. 2–6).
However, CN'885 does not explicitly disclose that before being introduced into the heat exchange tube of the quenching and condensation separation device 1, the SiCl₄ is preheated to 220–240 °C first, or that the SiCl₄ is introduced into the heat exchange tube of the quenching and condensation separation device 1 at a flow rate of 5–10 mL/min, as it describes running the fractional condensation zones generally by structural piping temperature controls without utilizing an injected SiCl₄ preheated fluid stream. (CN’885 Pgs. 2–7).
CN'816 discloses the missing silicon tetrachloride handling parameters and fluid introduction mechanics of the primary reference. CN'816 discloses a method for chloridizing roasting and metallurgical separation where a multi-stage fractional condensation system is used to handle gas-phase chloride mixtures. CN'816 explicitly teaches that silicon tetrachloride (SiCl₄) co-exists dynamically as a fluid stream alongside volatilized metal chlorides within these targeted gas-phase pathways. (CN’816 Pgs. 1, 3, 5–7).
CN'816 further teaches that the multi-stage condensation loop includes a distinct step where the roasting gas phase is selectively condensed across a first-stage temperature profile controlled at 80–260 °C to manage phase distribution. (CN’816 Pgs. 3, 5–8).
Preheating the SiCl₄ medium to a specific boundary of 220–240 °C represents a direct, predictable configuration to align the heat-exchange fluid with the broader 200–600 °C ferrous chloride capture window explicitly taught by the primary reference CN'885 to maintain strict local thermodynamic equilibrium, while running the fluid introduction line at a routine volumetric flow rate of 5–10 mL/min represents a standard scale adjustment to maintain stable counter-current cooling velocities inside the heat exchange tubes. (CN’885 Pgs. 2–6); (CN’816 Pgs. 3, 5–8).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to preheat the SiCl₄ fluid to 220–240 °C and introduce it into the heat exchange tubes of device 1 at a flow rate of 5–10 mL/min within the combined method. A person of ordinary skill in the art would have been motivated to preheat and regulate the delivery rate of the SiCl₄ stream because managing the temperature and mass flow of an injected cooling fluid is a routine engineering choice adjusted in gas-phase fractionation to maintain sharp condensation boundaries and prevent gas pipe fouling. (CN’885 Pgs. 2–6); (CN’816 Pgs. 1, 3, 5–8).
Combining the dual-condensation layout of CN'885 with the SiCl₄ gas metallurgy and condensation properties explicitly demonstrated in CN'816 represents a predictable application of known fluid mediums to tightly control the 200–600 °C fractional thermal zone, enabling the clean, steady-state capture of solid ferrous chloride fractions away from the downstream lithium residues. (CN’885 Pgs. 2–6); (CN’816 Pgs. 1, 3, 5–8).
As to Claim 10:
See the rejection of Claim 1, from which this claim depends, for a disclosure of the underlying recovery process steps; CN'885 discloses the method according to claim 1, wherein in step (5), the solid phase product obtained by the roasting is subjected to water leaching and stirring; and CN'885 discloses that the water leaching and the stirring are performed for 1–4 h. Specifically, CN'885 teaches taking the roasting slag and executing a water leaching step for an explicit duration of 1 hour or 3 hours to dissolve and separate the lithium component into the filtrate. (CN’885 Pgs. 2–6).
However, CN'885 does not explicitly disclose that the water used in the water leaching is deionized water, as it describes the leaching solvent generally as water or a water solution. (CN’885 Pgs. 2–7).
CN'328 discloses the missing deionized water leaching parameter of the primary reference. CN'328 discloses a method for recycling valuable metals from waste lithium-ion batteries where the processed battery mass is subjected to hydrometallurgical recovery. CN'328 explicitly teaches that the baked or roasted reduction powder is treated through a leaching step, specifically disclosing that the roasted product is soaked and leached using deionized water to selectively dissolve and isolate the target valence metal ions into a filtrate solution. (CN’328 Pgs. 1, 3–7).
Utilizing deionized water to execute the 1 to 4 h water leaching and stirring steps of CN'885 represents a predictable substitution of standard, high-purity industrial solvents entirely taught within the secondary art to prevent foreign ion contamination during the downstream recovery of battery materials. (CN’885 Pgs. 2–6); (CN’328 Pgs. 1, 3–7).
It would have been obvious to a person skilled in the art before the effective filing date of the instant application to utilize deionized water for the water leaching step of Claim 10 within the combined method. A person of ordinary skill in the art would have been motivated to select deionized water as the specific fluid medium for leaching and stirring the solid roasting product because eliminating elemental impurities and mineral variations from processing water is a routine variable optimized in hydrometallurgy to guarantee the high purity of downstream battery-grade precipitation products. (CN’885 Pgs. 2–6); (CN’328 Pgs. 1, 3–7).
Given that the primary reference CN'885 establishes an efficient timeline of 1 to 4 h for leaching its chlorinated solid slag, completing this step by substituting raw water with the deionized water explicitly demonstrated in the closely related battery-recycling stream of CN'328 constitutes a predictable application of known raw materials to ensure an uncontaminated aqueous matrix prior to precipitating lithium carbonate. (CN’885 Pgs. 2–6); (CN’328 Pgs. 1, 3–7).
Conclusion
The prior art made of record and not relied upon is considered pertinent to applicant's disclosure.
JP H086152 B2 discloses a method for recovering a noble metal such as platinum contained therein from a used electrode waste material of a fuel cell or the like.
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/JIMMY VO/
Primary Examiner
Art Unit 1723
/JIMMY VO/Primary Examiner, Art Unit 1723
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